Search Documents

By Gary Jobson

The Mt Todd Gold mine is situated near Katherine in the Northern Territory of Australia. Commissioning of an 8 Mtpa processing facility incorporating a four stage crushing circuit, ball mills, flotation, concentrate regrinding, high rate thickening and CIL processing was commenced during late 1996. A 4.5 M tonne /annum crushing and heap leach facility was commissioned during late 1994, and was subsequently upgraded to a nominal 6 M tonne / annum circuit during 1995. Processing equipment used included primary gyratory, secondary cone and tertiary Barmac VSI crushers. Some of the features of that crushing circuit and heap leach facility were presented at the Randol Perth 95 conference by myself. The location is described as sub tropical with average climatic conditions including 1100 millimetres annual rainfall, predominantly during the December. to March wet season, and 3000 millimetres evaporation and maximum temperatures ranging from around 26 to 46 °C. The main gold occurrence at the Mt Todd site is the Batman deposit, with current published Reserves of some 60 Mt@ 1.20 g Au / tonne ( 0.0425 oz Au / ton ). The primary ore from the Batman deposit is an extremely competent greywacke and is very hard at an average Bond Ball Mill Work Index of26.5 kWhrs/ tonne. Other indicators include a Bond Impact Crush Work Index of 20 kWhrs I tonne, an Abrasion Index of 0.15 and Ultimate Compressive Strength values of 100 - 200 MPa. The main gold occurrence is very finely disseminated free gold (at an average grain size of around 5 micron) within a combination of quartz and mixed sulphide. This gold bearing component is confined to largely parallel veins ranging from 0.5 to 25mm wide within a greywacke matrix. Although the gold is essentially non­refractory, there is a strong recovery versus grindsize relationship due mainly to the liberation characteristics and the relatively low permeability of the host greywacke rock.

Jan 1, 1997

By Richard Shea

Morris-Shea, an Alabama-based deep foundation contractor, is installing foundation piles at the sites of North America's two largest liquefied natural gas (LNG) terminals currently under construction. The contractor has crews working at Venture Global's LNG lique- faction plant in Plaquemines Parish, Louisiana, as well as constructing foundations for the Cheniere Energy Stage 3 Liquefaction Project in Corpus Christi, Texas. This work includes installing ap- proximately 27,000 DeWaal drilled, full displacement, cast-in-place concrete piles to support seven new midscale LNG trains at Cheniere's Texas ter- minal. Deep foundation installation began at Venture Global's Plaquemines facility, located approximately 30 mi (50 km) south of New Orleans, with an initial phase of 24,000 piles. Plans for the Plaquemines LNG terminal include three loading berths on the Mississippi River and four LNG containment tanks. The facility will utilize new pretreatment com- pressors, liquefaction trains and on- site gas turbine power plants to maintain an annual LNG export capacity of nearly 20 million tons (18 million tonnes) per annum. The site's LNG trains are configured in 18 blocks. The contractor is installing piles through an initial 6-20 ft (1.8-6 m) of cement mixed improved soils, which provide a safe work plat- form for equipment to operate on the extremely weak native soils. Drilling continues through soft clay in the subsurface and tips into a stiff clay or dense sand layer. Production piles are drilled to depths up to 124 ft (38 m) below ground surface. The firm is working with engi- neering, procurement and construction contractor KZJV, a Kellogg Brown & Root and Zachry joint venture, at the Louisi- ana location. The foundation team mobilized at the Plaquemines site in May 2022, to install 16, 18 and 24 in (400, 450 and 600 mm) diameter test piles to depths of 90 and 120 ft (27 and 36 m). The piling contractor returned in September 2022, to start Phase 1 of the deep foundation project, beginning installation of production piles.

Apr 1, 2023

By M. P. Amsden

"Ecstall Mining is a wholly-owned subsidiary of Texas Gulf Sulphur at Timmins, Ontario. The mill tonnage is 10,000 tons per day of two different ores, a copper zinc ore and a silver lead zinc ore, producing about 2400 tons of concentrates per day. These are as a copper concentrate, a high silver zinc concentrate, a low silver zinc concentrate and a silver-lead concentrate. For flotation recovery a very fine grind is required, about 80% minus 325 mesh and the concentrates produced are 40% minus 10 microns.The thickened concentrates are pumped to the filter floor for filtering and drying. The filtering and drying section is separated from the flotation section by a wall and is controlled from a dryer control room by one operator.Two operating panels are located in this control room. One contains stop-start controls for the thickeners, filter motors and vacuum pump motors, and stop-start controls for dryer feed conveyors and dryer product conveyors to concentrate load-out bins, as well as ammeters for conveyor drive motors. Dryer throughput is regulated by Merrick Type L weightometers located on the dryer discharge conveyors and recorded on tons per hour strip charts.The second panel contains the temperature strip charts, furnace drive ammeters, scrubber and exhaust gas controls, burner controls and panel alarms for the four dryers.Filtered concentrates are dried in Nichols-Herreshoff, multiple hearth, dryers consisting of nine hearths (zinc) and seven hearths (copper and lead). There are two dryers for handling zinc concentrates and one each for copper and lead concentrates. The dryers are heated by burning natural gas. Each zinc dryer handles 1,000 t.p.d. and the copper and lead dryers 750 t.p.d. each.Multiple hearth dryers are unusual in the base metal mining industry and the reasons for selecting them in preference to rotary dryers are as follows:1) The concentrates could only be filtered to 15 to 17 percent moisture. These sloppy filter cakes would be extremely difficult to handle in rotary dryers."

Jan 1, 1972

Have your name and address on the first page of your manuscript. Typewritten manuscript is preferred, but is not essential. If you have the manuscript typewritten, have it double spaced and written on one side of the page only; send two copies to us if convenient; only one copy of the illustrations is desired. In typewritten manuscript, formulas are better inserted by hand. Keep a copy of your manuscript in case of loss in the mail. ILLUSTRATIONS.-Photographs reproduce best when printed on glossy paper. Solio is one of the best varieties. A good print is preferable to a negative. Blueprints, drawings, or pen or pencil sketches can be used for line engravings, but care should be exercised to make them legible and clear, especially in the case of blueprints. Tracings or photographs of drawings are preferable to blueprints. Rough sketches can be used, if necessary. Photographs should be sent flat, not rolled. Do not put reference letters or other marks on photographs; if such letters or marks are necessary, fasten a piece of tracing paper to the top edge, trace a rough outline of the picture upon it and then place all letters, etc. on the tracing paper lithe exact position you wish them to appear. Restrict the number of micrographs as much as possible on account of the considerable expense involved. Standard magnifications for non-ferrous metals and alloys are recommended as follows: 10, 25, 50, 75, 100, 150, 250, 500, 750, 1000, 1500. Micrographs should be trimmed to 2 1/4 X 2 1/4 in. (six micrographs to a plate), or, if larger, up to 3 1/2 X 4 in. (two to a plate). The smaller size should be adopted if consistent with adequate representation of the subject. With each micrograph should be attached an explanatory title, together with the actual magnification, etching medium, treatment, etc., indicating what the author intends to show.

Jan 1, 1919

By E. Buehler, E. Berry

HE magnetic properties of dilute palladium alloys are of fundamental interest.''' Neutron diffraction studies3 have indicated that a better understanding of the behavior of these alloys could be realized with the availability of high-purity single crystals of palladium. Crucible melts of 99.999 pct pure palladium were found inadequate for such research. In a successful effort to fill this need, floating zone refining of palladium was undertaken. In this note the results of the initial experiments are reported. Zoning was accomplished in an argon atmosphere using 450 kc rf generator for an energy source. Samples 5 to 6 mm in diam were suspended in a gastight quartz envelope which could be moved vertically through the rf coil. A nominal upward traverse rate of 1.5 mm per min was maintained during a pass. Close coupling was found to be essential for stable retention of the liquid in the zone. Single crystals averaging 3.5 cm in length were obtained. Laue photography showed them to be continuous single crystals without gross imperfections. Several criteria were used for evaluation of the purity of the zoned material: magnetic behavior, low-temperature specific heat and residual resistivity ratios. Table I shows the residual resistance ratios obtained for two samples, one subjected to a single pass, the other zoned four times. It is appilrent that palladium can be zone refined with the electrically active impurities being transported toward the tail end. A decrease in susceptibility for zone-melted specimens over any previously measured material from 1.5" to 80'~ has been obserted by H. J. Williams and R. C. Sherwood4 substantiating the residual resistance data. The specific heat measurements are currently being made3 and will be published at some later date. Discussions with J. H. Wernick and S. C. Abrahams are acknowledged. We also wish to thank G. F. Brennert for the residual resistance measure ments.

Jan 1, 1962

By O’Brian. Daniel T.

Ore reserve determinations supporting property appraisals and economic analyses leading to investment decisions typically require many repetitive calculations. Computers have become well-accepted tools for accomplishing these computations, and are partly responsible for the development of new evaluation methods during the last decade. Excellent descriptions of manual ore reserve may be compared with current computer reserve methods1,2 in North America2 and South Africa3. A comparison indicates that computers are used in two cases: a) when a cost savings can be achieved by the simple conversion of a manual method to a computer processed equivalent, and b) when the quantity of data and ore reserve estimation purpose justifies adoption of a complex ore reserve computational method not feasible on a manual basis. Ore reserve computations essentially consist of using sample data and a mathematical method to arrive at an estimated geometry, tonnage, and grade for a mineral deposit. The mineral deposit is then subdivided, if necessary, into ore and waste components. This separation is based on economic factors which may be explicitly introduced in the computation or incorporated in a numerical cutoff grade4. Computer methods for accomplishing these tasks are presented in three papers. "The Use of Contour Surfaces as Predictive Models for Ore Values" discusses the requirements for a satisfactory value prediction model and the importance of considering auto -covariance. A weighted moving average method is described and compared to other mathematical surface development methods. In the second paper, "The Applications and Limitations of Computer Techniques in the Evaluation of Hypothermal Tin Lodes," particular problems encountered in estimating geometrically erratic economic deposits with a low valuable mineral content are documented. Economics and probabilistic concepts are introduced, allowing a comparison of risk and speculative value for a deposit. Economic considerations are integrated with ore estimation techniques in the third paper, "Method of Calculation of the Iron Ore Reserves in the Lorraine Deposit." The prediction model

Jan 1, 1969

By Heinz Boeni, Daniel Ott, Uca Silva

"Latin America is facing a rapid increase in internet use. Penetration with electronic equipment is in some countries approaching the level of industrialized countries. There is an evident need to resolve the management of “end-of-life” computers and other electronic equipment. Several studies in Latin America assessed the increasing e-waste quantities and confirmed the importance of a sustainable e-waste management. This paper gives an overview on the status on e-waste management in Latin America, explains the challenges for establishing an e-waste management system in a developing country setting and highlights the social and economic potential and the possibilities of a regional approach.IntroductionGlobal ICT-Trade and generation of e-waste from ICTThe global production of electronic devices and particularly of Information and Communication Technologies (ICT) faces the biggest industrial expansion of the history: OECD figures show that global trade of ICT technologies has reached 7.7% of the gross world product by 2004, the major proportion accruing from China [1]. In 2006 an estimated 230 million computers and 1 billion cell phones have been sold worldwide which corresponds to a volume of 5’848’000 t [2]. As a consequence, Waste from Electrical and Electronic Equipment (WEEE), or e-waste, is by far the fastest growing waste component. It reaches more than 5% referred to Municipal Solid Waste and e-waste generation in developing countries according to UNEP is expected to triple by 2010 [2].According to the OECD e-waste is “any appliance using an electric power supply that has reached its end-of-life” [3]. The European WEEE directive [4] distinguishes 10 categories of ewaste: Big household appliances, Small Household Appliances, IT and Telecommunications Equipment, Consumer Equipment, Lighting Equipment, Electrical and Electronic Tools, Toys, Leisure and Sports Equipment, Medical devices, Monitoring and Control Instruments, and Automatic dispensers. In this paper e-waste and WEEE are used as synonyms. Focus will be laid on e-waste from IT equipment (IT-waste), corresponding to category 3 of the European WEEEDirective."

Jan 1, 2008

By Georgeta Rata, Cornelia Petroman, Ioan Petroman

"Proper training and education in the field of recycling and waste treatment are inconceivable without the proper teaching of the terminology specific to both recycling (aerospace products recycling, automotive recycling, general recycling, recycling behaviour, recycling schemes, recycling systems, and waste recycling) and waste (characterization of waste, electronic waste, heavy-metal containing waste, monitoring of waste, radioactive waste, solid waste, waste collection, waste conversion, waste recycling, waste reutilization, waste stabilization, waste storage, and waste treatment) in both the students’ mother tongue and in English for Special Purposes (ESP) as a second language (with focus on the structuring of the specific terminology). Including linguistic approaches among the possible approaches of environmental protection and material sustainability means including linguistics among the various disciplines addressing energy supply, cement and building materials, metallurgy, chemistry, glassware, pulp and paper industry, machinery, automobile and electronic industries, collection, and sorting, further treatment and final disposal of post-consumer material.IntroductionCultural and technical changes have produced more novel artefacts than novel activities or novel properties, generating new vocabulary needs that are readily answered by new words (nouns).The purpose of this research is to show that ‘waste words’ – phrasal words (items that have the internal structure of phrases but function syntactically as words, e.g., cattle waste) and particularly compound nouns (words formed by combining roots, that tend to have a meaning that is more or less idiosyncratic or unpredictable, e.g., wastepaper ‘discarded paper’) – used in the study of the English of recycling, waste treatment, and clean technologies can be misleading for both native and non-native students and/or users of English because their structure alone does not allow instant, proper understanding of their meaning.Our hypothesis is that this inconvenient can be overcome by getting to reach the proper meaning of these new words. Or, arriving at the precise meanings of the ‘waste words’ depends on our knowledge of the world (for example, wastepaper indicates that some used paper is discarded and not wasted) rather than on purely linguistic knowledge."

Jan 1, 2008

By L. A. Morley, W. T. Parry

Engineering design problems encountered by mining engineers often depend on the properties of natural granular materials such as soil, poorly consolidated sediment, fault gouge, and hydrothermally produced clay. The physical properties of such granular materials frequently are determined by the properties of silicate clays. In most mining applications, the clays are in contact with water. Swelling of clay due to water infiltration can interfere with underground support and haulage systems. Water-saturated clay on fault, fracture, and bedding planes causes slope failures in open pit mines, roof falls and heaving of weak underclays in coal mines. Water may be removed and cementing agents may be produced in clays by electrical methods. When an electrical potential gradient is applied to a charged capillary by two electrodes, water moves toward the negative electrode.1 This electrokinetic phenomena is known as electroosmosis and has been used to successfully dewater clay minerals which lack strength because of water. A consolidation effect occurs when the flow of water is accompanied by a decrease in water content. Electrochemical changes in the clay result in cementing and irreversible hardening. Theory A mathematical basis of electroosmotic phenomena is presented by Helmholtz.3 Smoluchowski5 modifies Helmholtz’ model and presents Eq. 1 to describe fluid flow velocity due to electroosmosis. For this formula, the zeta potential can be measured experimentally. The Helmholtz-Smoluchowski model is reasonably valid for clay aggregates with large pores saturated with fresh water or dilute electrolyte solutions.° An assumption used for Eq. 1 (equations are given Table 1; terms are defined in Table 2) is that the thickness of the electrical double layer surrounding the clay particles is small compared with the pore radii. Because this is generally not the case, Schmid7 presents Eq. 2. The volume charge density in the pore may be approximated from cation exchange capacity and porosity measurements. Casagrande8 offers a semi-empirical relationship (Eq. 3) for volume rate of electroosmotic flow and finds that k, is relatively constant for most clays with an average

Jan 1, 1972

By Robert S. Mayo

My subject today is "Drilling-and-Blasting of Small Rock Tunnels". I am sure that there will be some people, primarily college students but perhaps Consulting Engineers, who will say: "Drilling-and-Blasting? I thought that went out with the invention of the Tunnel Boring Machine". A few years ago I talked to Dick Robbins and he told me that they did not try to sell a "Robbin's Mole" unless the tunnel was two miles long. For easy calculations I use 10,000-ft. (3km) A Mole today costs about $2 million which works out to $200 per lineal foot. This is "front money". You cannot place the order unless you are the uncontested low-bidder and then you must wait at least six months for delivery. Of course you can use that time to square up your portals, ready to turn-the-eye, or maybe you have shafts to sink to reach the tunnel grade. But this is "dead-time" and it is a heck of a feeling to have a job and cannot get started on it until you get the Mole. I do not have the American figures on lengths of tunnels but there was an article a few years ago in a British magazine: NEW CIVIL ENGINEER, on this subject. This stated that 75% of the jobs in Britain which were bid were for small tunnels, 10-ft x 10-ft (3m x 3m) or smaller, and that they were less than 5,000-ft (1.5km) in length. This does not mean that 75% of the footage in Britain was less than 10-ft in diameter. It means that 75% of the jobs being bid were less than a mile in length. From my experience in the U.S. I believe that we have substantially the same ratio: so many of the jobs with which I was associated were short, 1,000-ft, 2,000-ft and sometimes 3,000-ft and the bores

Jan 1, 1981

The mining industry has indeed gone global. International minerals producers are spending big money to move into countries that were off limits a few years ago. Due mostly to political reasons, these countries' huge mineral deposits - worth billions of dollars - were locked up and mining companies could only sit on the side and wonder "what if." Now much of the world is going the free market route. Once closed countries and developing nations are rewriting their mining and tax laws hoping to attract foreign investment to develop their natural resources. And big businesses - mining companies included - are lining up to get their share. But with new opportunities come new risks. While emerging nations are slowly accepting the free market concept and the privatization that goes along with it, many remain plagued with political instability and violence. Even developed nations carry their unique risks. Some of these include inadequate or contradictory legal and regulatory infrastructures, political patronage and outright corruption. So companies hoping to expand into new regions of the world must do more than examine the market. They must study the political stability and structure of a country, as well as its tax and environmental laws. And they must guarantee the safety of their employees. Control Risks Group is an international management consulting company that assists companies preparing to move into a new country. It employs about 225 people in 14 offices throughout the world. Control Risks personnel include lawyers, bankers and former law enforcement officers and intelligence people. The company held a one-day seminar in Washington, DC to update some of its clients on the state of affairs in various regions of the world. The seminar also covered extortion, corruption and product diversion. About 175 people attended the seminar. Speakers included Control Risks' operators, analysts and investigators. Here is a summary of what some of those people had to say.

Jan 1, 1998

By Harrison R. Cooper

With problems of diminishing ore grades and increasing cost of facilities, the mineral industry is acutely aware of the needs for improving beneficiation processes. Industry and universities are directing intensive technical effort to develop- ment of techniques with potential for using digital computers in automation and process analysis. The main problem of early attempts at on-line computer process control methods was that of economics. Now, with new advances in analysis and modeling making practical application economically important, it appears that minerals beneficiation is on the threshold of a significant jump in operating technology. In reviewing this work, particular note was taken of research and development applied to fundamental processes and the significant areas of effort where operations research techniques are used for process applications leading to the current state-of-the-art. Development of data automation and on-line control methods for process industries is deferred to the next section of this volume. This section deals with computer-facilitated mathematical-modeling development and reviews this work in conjunction with associated prior fundamental research leading to development of the methods. Flotation The 1960 state-of -the -art in flotation modeling was com- piled by Arbiter and Harris in the volume, 50th Anniversary of Froth Flotation, edited by Fuerstenau1. Many workers in the field of flotation initiated their research through flotation models, and this effort has led to use of simulation and operations re search methods employing computer techniques to facilitate advance of this work. The flotation model of Imaizumi and Inoue, published with supporting data in 1962, was expanded in 1968 with the objective of developing a basis for process automation.2,3 The fundamental work of Woodburn and Loveday in 1965 has led to techniques applicable to flotation circuit design optimization.4 This analysis is described in this volume by Woodburn in conjunction with King, Buchalter, and Piper. Niemi5,6 derived models for dynamics of a flotation circuit which were employed in developing a model describing

Jan 1, 1969

F. G. SMITH, Waterbury, Conn.-I would like to ask whether small amounts of iron give the maximum resistance at a low temperature, and if the large amounts of iron raise the temperature at which the maximum resistance occurs? That is, with a small amount of iron, will a maximum be at, say, 10°? F. E. BASH.-With the small amount of iron, the maximum is at a lower temperature, and with the increase of iron, the maximum increases; it seems to come to a maximum and then decrease; that is, increasing the iron up to a certain value increases the temperature at which there is a maximum of resistance, and any increase of iron beyond that value will start decreasing this temperature. With 2 per cent. of iron, the maximum is lower than for 1.5 per cent. which was the value that we found to be satisfactory. F. L. DRIVER,* Jr., Harrison, N. J.-I would like to ask what chance there is of the instrument companies getting together and deciding on standards for both manganin and constantan? We make both of these alloys in this country and we find that each one of the instrument com-panies who buy these materials require different standards of specific electrical resistance and thermal e.m.f. For instance, some of our cus-tomers require an electrical resistance in manganin or therlo of 240 ohms per circular mil foot and others require it with a specific electrical resist-ance of 270 ohms per circular mil foot. They buy these alloys in relatively small quantities so that it means much extra melting and supervision when making up small quantities for each consumer. If something could be. clone to eliminate this, it would be a benefit to the instrument com-panies, who could secure their material more cheaply, and to the manu-facturers, who would not have to spend so much time and effort in special manufacturing.

Jan 12, 1919

By M. J. Wilhelm

Bromine is the intermediate member of the halogen family of elements between iodine, a solid, and chlorine, a gas. The name is derived from the Greek bromos, meaning stench. Bromine is the only nonmetallic element that is liquid at room temperatures, with a melting point of -7.2°C and a boiling point of 58.78°C. Reddish brown in color, it dissolves easily in water to yield a red solution. Bromine volatilizes at ordinary temperatures to a red vapor that is very irritating to the eyes, nose, and throat. Density of the liquid is 3.12, and gas weighs 7.59 gpl in standard conditions. Because the element is highly active, it exists in nature only as bromide compounds. Bromine was discovered by Antoine Jerome Balard in seawater bitterns in 1825 and began to find commercial applications shortly thereafter in photography, medicine, and dyes. The first commercial production of bromine was in 1846 in Freeport, PA, followed by production in Germany in 1865. OCCURRENCE Bromine is widely distributed in nature, and compounds occur in solution in natural brines, in seawater, saline lakes, oil and gas well brines, and in evaporite chloride minerals. The concentration of bromine salts ranges from 60 to 65 ppm in seawater to several thousand ppm in oil well brines and saline lakes. The distribution of bromine in marine evaporite minerals and in the solutions from which they crystallized was determined many years ago by H.E. Boeke (1908). Theoretical aspects regarding the geochemistry of bromine have been published in numerous papers (Raup and Mite, 1969). Bromine occurs in solid solution in carnallite and sylvite associated in potash deposits, in halite in salt beds, salt in oil shales, and in trona deposits. Its presence is useful as a guide in geochemical exploration. The analysis of small amounts of bromides in the presence of large amounts of chlorides is discussed by W. Schwerdtner (1963). Manganese, iron, and iodine may interfere with the determination. Other analytical methods are discussed at length in literature (Stenger and Atchison, 1964).

Jan 1, 1994

By S. J. McCarroll

J. Bruce Clemmer, J. B. Rosenbaum, and C. H. Schack (U.S. Bureau of Mines, Salt Lake City)—We have watched with considerable interest Three Kids development of Manganese Inc. and have been impressed by the ingenuity with which management and staff have solved their many obstinate handling problems. Mr. McCarroll has been lenient in describing the difficulties faced by the company in developing a process for beneficiating Three Kids ore. The operation, equivalent in size to a 20,000-ton sulphide ore plant, is treating a unique ore by a flotation process pioneered by Manganese Inc. Perhaps the magnitude of the conditioning and flotation problems peculiar to Three Kids ore can be better appraised when it is pointed out that two-thirds to three-fourths of the ore is floated in the rougher step and that the four stages of cleaning which follow are equivalent to floating the whole plant feed two to three times. Our experience in both laboratory and pilot plant flotation of Three Kids and other manganese ores leads us to agree with Mr. McCarroll that suitable conditioning is a prime requisite for successful oil-emulsion flotation. The difficulties of proper conditioning were demonstrated in pilot plant processing of Artillery Peak manganese ore at our Boulder City, Nev., experiment station. Over three months of experiments and extensive plant revision were required before satisfactory conditioning was obtained in the 25-ton-per-day plant. Such factors as time, type and intensity of mixing, aeration, surface to volume relationship, and power input are still being studied by the Bureau of Mines in hopes of learning how best to condition manganese pulps for oil-emulsion flotation. Other factors affecting flotation of manganese with oil emulsion also deserve attention. A few of these are the influence of stability and proportion of ingredients in the collector emulsion, the feasibility of grinding

Jan 1, 1955

By Gene E. McClelland

4.1 INTRODUCTION Metallurgical testing is required on any ore, waste or tailing material before amenability to heap leaching treatment can be accurately determined. The amount of testing necessary to determine and optimize heap leach amenability and conditions for any feed material is determined by the ore type, grade, size of deposit, and the overall financial commitment a particular company is willing to make. It is very easy to either "under-test" or "over-test" a particular deposit or feed material. A good heap leach testing program will provide the necessary design information without "bogging" down in too much detail. The three basic stages of ore testing include: • Preliminary--bottle roll tests and small column percolation leach tests; • Detailed--larger column tests, different feed sizes, and feed size optimization; and • Pilot-Scale--Column or field heap. The following subchapters describe the stages of heap leach testing necessary to put a viable property into production. 4.2 PRELIMINARY METALLURGICAL TESTING The two types of preliminary tests that can be undertaken include bottle roll or cyanide agitation tests and small column leach tests. Each of these is described below, along with information about what the test results provide in terms of furthering knowledge of ore qualities. 4.2.1 Bottle Roll Tests The first step in a heap leach testing program is preliminary bottle roll tests at a relatively coarse feed size to determine the ore's leachability. Bottle roll tests are usually conducted on reverse circulation drill cuttings samples (= 1/4 inch (6 mm) in size). Bottle roll tests provide initial information about precious metal recovery, recovery rate, and reagent requirements. The recovery results are to be considered maximums because attrition grinding creates fines and liberates recoverable values that would not be liberated from the material during a static leach. Bottle roll tests can be conducted on up to 25 lb (11 kg) ore charges of up to two inches (50 mm) in size. However, the coarser the feed the more attrition grinding occurs during rolling.

Jan 1, 1988

By R. L. Frink

I FEAR that my treatment of this subject may not, in all instances, meet the approval of those who read my opinion as to the utility and efficiency of pyrometers in the making of glass, or bottle-glass. It may be superfluous for me to add that this opening statement is based on over 15 years' experience in an endeavor to successfully apply pyrometers, or heat-measuring instruments, to glass-melting furnaces, particularly the type known as tank-furnaces, and that such endeavors have proved more or less futile. It is my desire, therefore, to herein set forth the problems encountered, hoping thereby to stimulate further effort in the successful application of pyrometers. It is my desire, also, to invite criticism of the methods used and suggestions from those who may have a wider experience, or who may have better methods. Fin. 1.-GLASS TANK FURNACE. In general, I believe that the inefficient application of pyrometry to the melting of glass in tank furnaces is not due as much to the instruments as to the conditions under which they must be used and the character of the medium in which they must reside. Most of those present no doubt, are familiar, at least in a general way, with the methods and apparatus used in the melting and making of bottle-glass, or glass that is formed into containers of various shapes and designs. Most of this glass is made in what is known as a tank furnace, these furnaces vary widely in design and construction. In general, they consist of a rectangular tank-like form, built up of fireclay blocks, and range from 2 by 4 ft. (0.6 by 1.2 m.) to 24 by 140 ft. (7.3 by 42.6 m.) with a depth of glass from 2 to 5 ft.; or a capacity of from approximately. 2 to 700 tons. Fig. 1 shows the general

Jan 9, 1919

By W. J. Shearer

A COMPARISON of statistics relating to coal min-ing in Great Britain and the United States leads one to suspect widely differing conditions and practice. A brief review of some of the conditions may enable American mining men to appreciate more fully the difficulties with which British engineers are faced, and to some extent explain why American, coal can be sold f.o.b. mine at the present time at about 50 to 60 per cent of the cost of British coal at the mine, and why British man-shift outputs fall far below American at present being less than 30 per cent. Bulman gives the average depth of 6-ft. seams. in Britain at 1350 ft., and of 2-ft. seams at 517 ft. George S. Rice states that 90 per cent of American coal 'is taken from less than 600 ft. Much English coal is taken below 2000 ft. and some below 3000 ft. The total influence of depth on cost is not always evident. Room work as known in America becomes less advisable with increasing depth, as pillars crush badly if left -for any length of time. A limited amount of bord and pillar work, taking very small amounts of coal in the first working, takes the place of room work, with rapid working of the pillars as soon as the panels are broken. Ventilation difficulties increase tremendously with depth and thinning coal. In America the chief use of air underground is to dilute gas. In Britain it serves a dual purpose, in many cases the amount of air cir-culated being far in excess of the quantity necessary to dilute gas to reasonable proportions. At great .depths rock temperatures exceed-95 deg. F., and great volumes of air are necessary for rock cooling. At these high temperatures humidity becomes an important factor. At present a great deal of research is under way to find methods of overcoming the troubles experienced with high temperature and humidity in deep mines.

Jan 8, 1928

By J. F. Crouch

INTRODUCTION Roll front uranium has unique geologic characteristics (L,) which require specialized mining procedures. Roll front uranium occurs in arkosic sandstones with the ore zone typically less than 100 feet in thickness with up to 400 feet of overburden. The areal extent of ore is often a few thousand acres. Within the ore zone, ore grade material occurs in C and S shaped rolls, which are a few feet in width and height but may run laterally for several hundred feet. Since the total mine plan may include mining all ore, a roll front uranium mine may cover a few thousand acres. For both economic and environmental reasons, a roll front deposit is mined as a series of phases or sub-pits (3, 4) as illustrated in Figure 1. A typical sub-pit is 400 to 500 feet across. Sub-pit 1 is stripped initially and the overburden is spoiled on the surface. Depending on haul distance and overburden characteristics, stripping is done with trucks and shovels (or front-end loaders) or scrapers. When ore is encountered in sub-pit 1, mining begins, and the stripping equipment is moved to sub-pit 2, with sub-pit 2 overburden also being spoiled on the surface. When ore is encountered in sub- pit 2, sub-pit 1 mining should ideally be complete so the mining equipment can move to sub-pit 2 and sub-pit 3 overburden can be spoiled in the mined out sub-pit 1 area. Ore is mined selectively in a series of 12 to 24 inch cuts by a 2-1/2 yard backhoe on the basis of radiometric probing. For effective drainage, ditches must be maintained on the boundary of sub-pits, and mining lifts are kept as level as possible, with the orebody slightly elevated as part of general good mining practice. Waste removal in the ore zone is done by self- loading scrapers near the orebodies and by push-pull or push loaded scrapers away from the orebodies. Each self-loading scraper cut is

Jan 1, 1983

By Farouk Tedjar

The growth of miniaturization in one hand and energy density need lead to rapid development of new chemistry for batteries as observed during last decade particularly in secondary systems. However dark side of the progress is always in environment impact and scarcity of resources. In the lithium segment a jump from six electrochemical series (all non rechargeable systems) before 1990 to more than 15 systems including new rechargeable systems has been observe during last decade. Lithium battery chemistry was enriched with new electrolytes, conductive salts and new oxide composition. In the same time, lithium ion battery safety was strongly improved during last period. The progress of these technologies was strongly boosted by the substantial growth of the mobile telephone market in the mid 1990's (+ 1% a month in 2000, and a 71% penetration rate in Europe in 2005 compared to just a few percent in 1996). This market is constantly in search of higher electrochemical performance (mainly energy density in front of increasing miniaturization. Recycling of end of life product could be managed not only in environment aspect, but also under resource conservation and economic aspect of arising technology (mainly on the effect of re-injection recovered materials) The Li-ion segment is a new branch in the cobalt market and impact of LiCoOx will be fitted with the actual increase of the demand for this battery. For this reason all the processes aim to valorize active and packaging material. RECUPYL process project was conducted in this direction to offer a new solution for recycling those lithium systems. The process has been developed at pilot industrial scales using synergy between new equipment (for mechanical pre-treatment step) and existing line at RECUPYL facility (for chemical treatment) and should lead to high valorization rate with a minimum hazardous co-products. The process was also developed in order to be able to recover news cathode materials introduced in electrode like LiFEPO4, LiMxCoyO2 and new anode composition in substitution with former carbon anode.

Jan 1, 2006