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By J. E. Dutrizac, Alba Sunyer

The factors affecting the precipitation of arsenic-free alunite (KAl3(SO4)2(OH)6) were investigated in a series of laboratory experiments. Although the amount of alunite precipitated increased with increasing retention time, temperature, solution pH and the concentrations of Al(SO4)1.5 or K2SO4, the composition of the products remained nearly constant. The preferred conditions identified for the precipitation of arsenic-free alunite were then used to assess the behaviour As(III) and As(V) during alunite precipitation at 98°C. Increasing additions of As(III), as NaAsO2 or As2O3, resulted in a slight increase in the amount of precipitate formed at pH 2.80. However, X-ray diffraction analysis showed all the precipitates to be alunite, and <0.2% As(III) was detected in the products. In contrast, increasing additions of As(V), as KH2AsO4 or As2O5, resulted in significant increases in the amount of precipitate formed at pH 2.80 or 3.00. Furthermore, the arsenate (AsO43-) contents of the precipitates increased up to 15%, and the sulphate contents decreased, as the As(V) concentration of the synthesis solution increased from 0 to 9 g/L. X-ray diffraction analysis of the products detected only alunite, but there was a progressive expansion of the alunite cell with increasing arsenate content. The implication is that arsenate substitutes for sulphate in the alunite structure. Preliminary observations on the environmental stability of the arsenate-bearing alunite using the TCLP protocol gave solubilities <0.4 mg/L from a precipitate containing 8.5% AsO4 and 7.8 mg/L from a product containing 19.5% AsO4. Thus, it appears that alunites containing less than -10% AsO4 would be an acceptable vehicle for the disposal of arsenic from hydrometallurgical processing solutions.

Jan 1, 2011

By F. D. Richardson

STAFF: Editor, Gerhard Derge Carnegie lnstitute of Technology Schenley Park Pittsburgh, Pa. 15213 Editorial Assistant, M. A. Redmerski Production Editor, Otto T. Johnson THE METALLURGICAL SOCIETY: Jack H. Scoff, President; Richard C. Cole, Past-President; Harold 0. Emerick, President-Elect; Ralph L. Hennebach, Treasurer; Robert W. Shearman, Secretary; Donald A. Parks, Assistant Secretary; Leslie S. Wilcoxson, Assistant Secretary. PUBLICATIONS COMMITTEE OF THE METALLURGICAL SOCIETY: R. E. Lund, Chairman; T. E. Dancy, H. O. Emerick, P. R. Gouwens, R. E. Grace, J. H. Jackson, H. H. Johnson, A. E. Lee, Jr., R. R. Nosh, R. D. Pehlke, and Paul Queneau. Review and selection of manuscripts is under the supervision of the following Publications Committees: Extroctive Metallurgy Division: R. E. Grace, Choirmon; G. M. Bell, D. W. Bridges, J. M. Cigon, G. W. Heoly, T. A. Henrie, R. D. (Holli-doy, C. E. O&apos;Neill, C. S. Somis, A. W. Schlech-ten, and Charles Shull. Iron and Steel Division: R. D. Pehlke, Choirmon; T. P. Floridis, Arnulf Muon, G. R. St. Pierre, E. T. Turkdogan, L. H.Von Vlock, and J. M. Uys. lnstitute of Metals Division: H. H. Johnson, Jr., Chairman; J. B. Clark, Vice Choirmon; M. R. Achter, G. F. Bolling. J. O. Brittoin, G. P. Conard, II, R. S. Davis, L. A. Girifolco, G. T. Hohn, J. P. Hirth, L. Himmel, G.T.horn, George Krauss, J. D. Livingston, R. D. Manning, H. M. Otte, S. V. Rodcliffe, R. E. Reed-Hill, H. C. Rogers, E. ScaIa, O. Sherby, M. T. Simnad, G. R. Speich, E. A. Steigerwald, R. A. Swolin, W. A. Tiller, W. R. Upthegrove, J. B. Wagner, Jr., Derek Wolton, H. W. Weart, W. W. Webb, and D. Published seven times yearly in February, March, April, June, August, October, and December by the American lnstitute of Mining, Metallurgical, and Petroleum Engineers, Inc., 345 East 47th Street New York, N. Y. 10017. Telephone PLAZA 2 -6800. Subscription $30 per year for non-AtME members m United States and North, South, and Central America; $32, foreign; $7.50 for AlME members Single copies, $6.00, single copier foreign, $6.50. The AiME is not responsible for any statement made or opinion expressed in its publications . . . Copyright 1964 by the American Institute of Mining, Metallurgical, and Petroleum Engineers, Inc. . . . Registered cable address, AiME New York Indexed in Engineering Index, Industrial Arts Index, and Chemical Abstracts. Second dost postaas poid ot New York, N. Y.,ond Ann Arbor, Mich. Volume 230, Number 6.

Jan 1, 1964

By Steven Gillgrass

Gerdau Monroe has installed and commissioned the Tallman Supersonic Carbon Injection system (TSCi™) for use in their 135-ton Electric Arc Furnace (EAF) located in Monroe, MI. TSCi™ is an innovative injection system with the primary aim of improving injection carbon delivery to the steel bath by increasing efficiency. It is proven technology with installations at numerous steelmaking facilities all over the world. Results from installation and commissioning of TSCi™ at Gerdau Monroe will be reported as well as on-going trials and optimization efforts. Standard injection systems deliver carbonaceous material from a vessel to the EAF using air where it is injected through a lance pipe or oxy-fuel burner. Using this method, material delivery to the EAF is limited by the pressure and velocity of the transport air. Loss of efficiency can be observed by looking at the quality of the injection carbon stream during operation or by measuring loss on ignition of EAF slag or off-gas system dust. TSCi™ uses a patented design and proprietary technology to improve delivery of carbonaceous material to the EAF. Using this method, savings at Gerdau Monroe were realized by reducing the amount of carbon used in each heat while maintaining or improving upon existing foamy slag practice. As well, significant savings were achieved as a result of reduced FeO content in the slag and improved recovery of iron. Additional savings can be realized by using a smaller mesh material to take advantage of lower material cost and higher injected surface area. Other inherent benefits to improved injection carbon include lower energy and electrode consumption, reduced refractory wear and reduced maintenance cost.

Oct 30, 2017

By Steven Dapp

A new hyperbolic model for generating P-Y curves is developed for use in lateral analyses of cemented soils (a ?C-F? soil with both cohesive intercept and friction angle), and includes degradation of the static loading soil response to model the effects of load cycling (repeated loading) with load cycle number. The p-y curves, for both static and cyclic loading, were derived from full-scale lateral test results conducted on a set of six drilled shafts in a loess deposit at the Kansas DOT test site. Values of the model parameters obtained from the load testing at this site are presented, as well as a discussion of the model parameters effect on the results such that the model may be rationally applied to other sites with C-F type soils. A step-by-step procedure is presented to illustrate the application of the model. Correlation of the model to the CPT cone tip resistance (qc) from CPT testing is made, as this was found to be the most reliable; However, the model is applicable to other methods, both in-situ and laboratory, of assessing the soils strength.

Jan 1, 2011

By M Dickie

During the 1980s and early 1990s low cost gold treatment plants were at the forefront of the Australian engineering and mineral processing business. This opportunity was the result of a high gold price and innovative design by emerging engineering companies in an expanding market. The low cost approach to gold plant design and project execution resulted in smaller and shorter life projects becoming viable and the multiplicity and fast track approach of the projects resulted in sharp learning curves. From the mid 1990s to the early 2000s the relatively low commodity prices and resultant reduction in minerals opportunities also resulted in innovative design in both small and large plants, albeit on a reduced scale. The innovation was associated with alleviating capital cost pressures for large capacity low-grade (eg NewcrestÆs Cadia) or small-scale (eg JubileeÆs Cosmos) projects to either improve NPV or to enable the project to be won and profitable in the competitive lump sum environment. Between 2004 and 2008 the engineering market became less competitive from an engineerÆs perspective due to the high demand for engineering services in a high price commodity market. The need for innovation remained but was driven overtly by shortages in water, power and personnel and long equipment lead times. This paper discusses some of the factors that contribute to the capital cost of a project and the scope for improving the likelihood of financing and increasing the project net present value (NPV). Case studies with recent large copper concentrator projects will be used to demonstrate how a value driven approach to project and plant layout can result in significant improvement in NPV.

Jan 1, 2009

By C. Musingwini

The Bushveld Complex is economically significant and strategically important to South Africa, thus it is imperative that optimal extraction of platinum group metal (PGM) resources on the Bushveld Complex is achieved. Optimal extraction broadly requires that the maximum amount of ore is extracted by excavating and hauling the minimum amount of waste in the shortest possible time, at the least cost, and in the safest and most environmentally acceptable manner. In open-pit mine planning this entails among other things, minimizing the waste stripping ratio, whereas in underground mine planning it includes minimizing the metres of waste development. In conventional mining, the main development that is in waste or partly in waste and defines the mining grid pattern includes levels and raises. It was prudent to consider optimizing level and raise spacing in conventional mining because the method is a prevalent mining method on the Bushveld Complex accounting for nearly 70% of platinum production.

Jan 1, 2010

By François G. Bernardeau

Cement-bentonite (C-B) slurry walls have been used for ground water cutoff in deep excavations, contamination containment, and other applications. Hydraulic conductivity is of great concern for cement-bentonite slurry walls. The quality control program usually specifies permeability and strength tests to be performed in a laboratory on samples cast in the field and cured for a certain time. Currently there is no widely accepted in-situ test method for the performance of the constructed slurry wall. A recently constructed cement-bentonite slurry wall in South Carolina included a 200-foot-long test section done in advance of the production of an 8,300 foot-long wall. Intensive tests and investigation were conducted during and after the test section construction: Monitoring piezometers were installed using wet insertion method and packer tests were conducted in boreholes drilled in the post-28-day-cure slurry wall. The permeability measured from the packer tests was about two orders of magnitude higher than the design and the laboratory test values. Additional investigation was performed using a video camera to inspect the borehole walls especially the packer test sections. Test pits were excavated to observe both sides of the cement-bentonite slurry wall. A thorough inspection was conducted to evaluate the integrity of the wall. The lessons learned and issues regarding the disturbance effects of in-situ testing of cement bentonite slurry walls is presented and discussed.

Jan 1, 2008

By D Craw, N Newman

Previously published stream quality data and AMD (acid mine drainage) analyses have been collated into a database (DAME û Database for Assessment of Mine Environments) and additional samples including NMD (neutral mine drainage) have been collected from coal mine drainages on the West Coast. Variations in the chemistry of mine drainage samples can be related to four factors: 1.&nbsp;&nbsp;&nbsp;&nbsp;&nbsp;&nbsp; regional geology (Brunner Coal Measures or Paparoa Coal Measures) 2.&nbsp; &nbsp;&nbsp;&nbsp;&nbsp; mine type (open cast or underground) 3.&nbsp;&nbsp;&nbsp;&nbsp;&nbsp;&nbsp; hydrogeology 4.&nbsp;&nbsp;&nbsp;&nbsp;&nbsp;&nbsp; local geology &nbsp; In general mines hosted in Paparoa Coal Measures produce NMD whereas mines hosted in Brunner Coal Measures produce AMD with highly variable chemistry. Differences in coal mine drainage chemistry between mines hosted in Brunner and Paparoa Coal Measure sequences can be measured with acid-base accounting geochemistry. Correlation of ANC (acid neutralising capacity) and MPA (maximum potential acidity) analysis of rocks with mine drainage samples is substantially better than correlation of NAG (net acid generation) analysis with mine drainage chemistry. &nbsp; It is likely that NAG analysis over estimates acid generation potential of some samples due to release of organic acids that are not released under environmental conditions. In general, differences in mine drainage chemistry between Brunner and Paparoa hosted mines are attributed to the effect depositional environment and diagenesis on coal measure composition. Exceptions to generalisations relating mine drainage to Paparoa or Brunner host rocks are likely where local geological conditions cause increased deposition of pyrite in Paparoa rocks and additional analyses are required to confirm interpretations. &nbsp; Based on the interpretation of these results a mine drainage chemistry hazard model has been constructed. The hazard model predicts likely variations in mine drainage chemistry based on the four factors listed above. The hazard model has implications for mine related water quality risks. For example, ecotoxicity, human health affects and selection of remediation techniques are all sensitive to mine drainage chemistry.

Jan 1, 2006

By S. Chung

Mining in highly stressed ground conditions has been an increasing challenge at the Brunswick mine. Over the past 30 years, two large regional pillars were established on the 1000 m Level to support the ground for safe extraction of the surrounding ore. The South Regional Pillar contains about 5 Mt of ore and the North Regional Pillar has about 3 Mt of ore. In the fall of 2000, mining activities were suspended following a series of ground movements around the South Regional Pillar along weak geological structures. After extensive geotechnical investigations, a decision was made to carry out a mass de-stress blast in the West Ore Zone (WOZ) to cut the principal stress across an area of 135 m along strike and 86 m high. At the end of July 2001, the WOZ mass blast was successfully initiated. Following this success, the mine decided to take a similar approach to de-stressing the North Regional Pillar (NRP). The final mass blast was taken in early February 2003. Advanced blasting technologies, including state-of-the-art electronic detonators and blast simulation, played an important role in achieving the successful blasts. The high precision of delay timing and long delay duration of the electronic detonators ensured accurate and reliable initiations of over 800 blastholes in the WOZ mass blast. Orica’s DMC_Blast (Distinct Motion Code) simulation code was used to study different blasting blocks for the desired timing in the initiation sequence. This paper describes the process of applying electronic detonators in complex blasts, and discusses how blast simulation results were interpreted to guide engineering design of the mass blasts.

Jan 1, 2005

By B. H. Bergstrom

The Gaudin-Meloy1 size distribution equation requires the evaluation of two constants, xo and r. The common log-log representation of a screen analysis as cumulative weight percentage passing versus size, readily yields the required constants as illustrated for curves A and B in Fig. 1. The 100% passing (y = 1.0) size is simply x,. If the straight portion of the curve be extrapolated to the 100% passing (y=1.0) abscissa, the intercept is the Schuhmann size modulus, k. The ratio of xo to the size modulus, k, is r as shown: The Gaudin-Meloy equation requires that the slope, a, of the size distribution approach unity in the finer sizes, as is the case for curves A and B. Often distributions are better characterized by slopes other than unity. The schuhmann2 equation is able to represent these distributions if the upper portions of the curves are ignored. Empirically, the Gaudin-Meloy equation can be modified to yield slopes other than unity by writing it as Here, a is the familiar Schuhmann slope, and xo is the 100% passing size. The value of r is still given by Eq. 2. Eq. 4 with its easily determined constants is more versatile than either the Schuhmann or the Gaudin-Meloy equations, since a wide variety of size distributions can be closely described. It is easier to fit than the Rosin-Rammler3 equation, and has the advantage that the top size, x,, is finite. REFERENCES A. M. Gaudin and T. P. Meloy: Model and a Comminution Distribution Equation for Single Fracture, AIME Transactions, 1962, Vol. 223, pp. 40 to 43. R. Schumann, Jr.: principles of comminution, I., size Distribution and Surface Calculations, AIME Technical Publication No. 1187, 1940. P. Rosin and E. Rammler: The Laws coverning the Fineness of Powdered Coal, J. Inst. Fuel, 1933. Vol. 7, p. 29.

Jan 1, 1967

By W. A. Heitbrink, J. D. Ashley, R. E. Gresh, J. A. Organiscak, A. B. Cecala

Many different types of mobile equipment used in surface coal mining utilize enclosed cabs to protect equipment operators. The overburden removal process is extremely dusty and can cause excessive exposure to respirable dust, especially crystalline silica. After equipment is used for years, many components of the enclosure deteriorate and its effectiveness is greatly reduced. This report discusses a cooperative research study performed on an Ingersoll Rand DM45E surface drill, retrofitted with a new Sigma pressurization and filtration system. Respirable dust concentrations were substantially reduced from 0.64 mg/m3 during pre-testing to 0.05 mg/m3 during post-testing with the new system, representing a 92% reduction in dust levels in the drill cab. This new system appears to be a very well-built and sturdy device, well-suited for the mining industry.

Jan 1, 2002

By J. A. Organiscak

Water spray applications are one of the principal means of controlling airborne respirable dust in coal mines. Since many coals are hydrophobic and not easily wetted by water, wetting agents can be added to the spray water in an effort to improve coal wetting and assist with dust capture. In order to study wetting agent effects on coal dust capture, laboratory experiments were conducted with three wetting agents used by the coal industry on -325 mesh sized Pocahontas No. 3 coal dust. Significant differences in coal dust sinktimes were observed among the three wetting agents at water mixture concentrations of 0.05%, 0.1%, and 0.2%. The best wetting agent as identified by the coal dust sink test was only tested at the lowest 0.05% water mixture concentration and was found to have a negligible effect on spray airborne dust capture. Water spray airborne dust capture results for all three wetting agents tested at a 0.2% water mixture concentration showed that all three wetting agents exhibit similar but small improvements in dust capture efficiency as compared with water. These results indicate that the coal dust sink test may not be a good predictor for the capture of airborne dust. Additional research is needed to examine if the coal dust sink test is a better predictor of wetting agent dust suppression effects during cutting, loading, conveying, and dumping of coal products by comparison to airborne dust capture from sprays.

Feb 27, 2013

By M. King Hubbert

In 1856 Henry Darcy described in an appendix to his book, Les Fontaines Publiques de la Ville de Dijon, a series of experiments on the downward flow of water through filter sands, whereby it was established that the rate of flow is given by the equation: q= - K(h2- h2)/I1 in which q is the volume of water crossing unit area in unit time, I is the thickness of the sand, h1 and h2 the heights above a reference level of the water in manometers terminated above and below the sand, respectively, and K a factor of proportionality. This relationship, appropriately, soon became known as Darcy&apos;s law. Subsequently many separate attempts have been made to give Darcy&apos;s empirical expression a more general physical fornzulation, with the result that .so many mutually inconsistent expressions of what is purported to he Darcy&apos;s law have appeared in published literature that sight has often been lost of Darcy&apos;s own work and of its significance. In the present paper, therefore, it shall be our purpose to reinform ourselves upon what Darcy himself did, and then to determine the meaning of his results when expressed explicitly in terms of the pertinent physical variables involved. This will be done first by the empirical method used by Darcy himself, and then by direct derivation from the Navier-Stokes equation of motion of viscous fluids. We find in this manner that: q = (N4(p/p)[g - (l/p) gradPI = WE, is a physical expression for Darcy&apos;s law, which is valid for liquids generally, and for gases at pressures higher than about 20 atmospheres. Here N is a shape factor and d a characteristic length of the pore structure of the solid, p and IL are the density and viscosity of the fluid, a = (Nd2)(pip) is the volume conductivity of the system, and E = [g - (l/p) grad p] is the impelling force per unit mass acting upon the fluid. It is found also that Darcy&apos;s law is valid only for flow velocities such that the inertial forces are negligible as compared with those arising from viscosity. In general, three-dimensional space there exist two superposed physical fields: a field of force of charucteristic vector E, and a field of flow of vector q. The force field is more general than the flow field since it has values in all space capable of being occupied by the fluid. So long as the fluid density is constant or is a function of the pressure only, curl E = 0, E = - grad F where F = gz + dp/p The field of flow, independently of the force field, must satisfy the conservation of mom, leading to the equation of continuity div pq = — f ap/at , where f is the porosity, und t is the time. For steady rnotion ap/at = 0. and div pq = 0 .

Jan 1, 1957

By Richard C. Moore

The U.S. Bureau of Mines contracted a study to review the air and water monitoring requirements of five Federal environmental acts (the Federal Water Pollution Control Act, the Safe Drinking Water Act, the Clean Air Act, the Surface Mining Control and Reclamation Act, and the Resource Conservation and Recovery Act) and to compile information on the currently available monitoring instruments. A literature review of the acts and their associated regulations was conducted to determine applicability to the mining industry as well as any air and water monitoring requirements. In addition, an industry search and literature review was performed to assess the instrumentation available to comply with the monitoring regulations. Based on this research, the instruments are classified by general applicability to air, surface water, or ground water monitoring, then catalogued by monitoring parameters. A discussion is provided on recommended and innovative monitoring techniques.

Jan 1, 1981

By E. D. Gardner

This is one of a series of papers describing mining and related subjects affecting mining in western mining districts and mineralized areas. The parts of this paper pertaining to current production, mining and milling methods and practices, and general conditions affecting mining were collected principally during a field survey made in May 1935. Some original data obtained on previous visits to the area are also included. The history, geological background, and past production of the mines and district are largely abstracted from previous publications, for which due credit is given later in the test. The principal districts in the Black Mountains are the Oatman or San Francisco and the Union Pass or Katherine. This range contains the principal producing mines of the State, in which gold is the only important metal. Although the total production of gold is relatively small as compared to other districts where the precious metals are obtained as a by-product in copper min¬ing, the area is one of considerable economic importance to the State.

Jan 1, 1936

By D. R. Dolinar

The fall of small pieces of rock between roof bolts causes over 80 pct of the ground fall injuries in underground coal mines. Roof screen, if installed on the bolting cycle, can significantly reduce the number of these injuries. As the surface of the roof fails, the screen will have to retain the broken rock without failing or excessive deformation. Welded wire screen is the most common type of screen material used for roof surface control in underground coal mines. The load-deformation characteristics and the load capacity of welded wire screen were evaluated in the laboratory in a specially built test frame. In the frame, the square screen sections were held in place by bolts and bearing plates at the four corners while a center load was applied to the screen. The parameters evaluated included the wire gauge and configuration, the bearing plate loads and bolt spacing. The wire gauge of the tested screen was 4, 6, 8, 9 and 10. Maximum screen capacities for both load and displacement were developed for each of the wire gauge sizes along with the screen stiffness. This information can be used for screen selection and for the evaluation and determination of screen performance. DISCLAIMER: The findings and conclusions in this report have not been formally disseminated by the National Institute for Occupational Safety and Health and should not be construed to represent any agency determination or policy.

Jan 1, 2009

By L. F. Bishop

THE ore bodies of the Butte district1 are found in many different vein systems having many different structural characteristics; some are narrow with self-supporting ore but with weak walls; some are wide with weak or strong ore and weak walls; some of the veins dip steeply and others do not; some have well defined walls, and others, in the "horsetail" area, are mined to sample limits. Consequently, many types of stoping methods are used in mining these ores. The stoping methods used in the Anaconda Copper Mining Company&apos;s properties are the square-set and cut-and-fill. The square-set method is the principal stoping method used because, generally speaking, the ore and walls are weak, the ore is high grade, and good extraction is required. These square-set stopes may be classified as conventional, vertical slice, slot, timber rill, and Mitchell slice types. The conventional type square-sets are those of various shapes and sizes, mined by horizontal slicing to the boundaries of the stopes. The rock is slid to the chute and the filling is moved by gravity. The ore and waste may be handled with a scraper when local conditions require it. The vertical-slice square-set system is the type used in narrow veins that are mined in vertical panels. The slot stopes are stopes in wide veins that employ a slot across the vein instead of a con- ventional chute to handle the ore. The timber rill fs a type of square set in which the ore is mined as in inclined cut and fill but the square-set timber is used. The Mitchell slice is the type used for some broken or heavy ore blocks, or for pillar extraction between two converging slot stopes. The cut-and-fill method is used where the ore will stand without much support although the walls may be weak. The principal type of this method is the horizontal cut-and-fill, with mechanical scraping. The inclined cut-and-fill, or rill, stopes are used in some veins too narrow for scraping. The production from the various meth: ods in use for the second six months of 1944 are shown in Table I. [ ] This paper deals mainly with the vertical slice and slot types of square-set stoping, because of their effect on production locally and their potential possibilities here and in other districts, with sufficient reference to conventional square-setting for comparison.

Jan 1, 1945

By G. R. Addie

"In February, 1975, the President of Georgia Iron Works Company, Mr. Tom Hagler decided to build a new hydraulic laboratory to water performance test all of the companies' centrifugal slurry pumps.Equipment procurement started in March, 1975, by ordering of a mini computer data acquisition system.Work commenced on the lab in July, 1976, by erection of a Butler Building and the laying down of a 20,000 gallon concrete sump.By late October, a facility compromising two water test loops of 8"" and 18"" pipe diameter was in operation carrying out head quantity and suction performance tests on pumps up to 16"" in size.In August of the following year, an 8"" diameter pipe slurry loop was installed to enable pipe friction and deposition velocity of a slurry to be measured in an 8"" inside diameter pipe.With the encouragement of Reserve Mining in Minnesota and IMCC in Florida, this was followed in March, 1978, by the addition of over 300 feet of 18"" diameter pipe to make up an 18"" diameter pipe slurry system complete with head loss and deposition velocity observation sections.The office and shop areas of the lab were again expanded in April, 1978, to provide further expansion to the test sections and staff facilities."

Jan 1, 1980

By Hendrick A. D. Kirsten, A. Clayr Alexander

Stillwater Mining Company is currently developing a second underground palladium and platinum mine approximately 21 km (13 miles) west of their existing Nye operation in the Bear tooth Mountains of southern Montana. Portals for the two access tunnels required at the new site were excavated in limestone talus to accommodate4.6 m (15 feet) diameter tunnel boring machines. A review of the technical design, field conditions and construction of the portals is presented.

Jan 1, 2001

By Pekka Taskinen

Advanced optimizing control of the fluidised bed zinc roaster has been in use for an extended period at Boliden Kokkola and the implications are seen as extended bed stability, increased throughput and far better on-line availability. The roaster furnace control system developed is based on a dynamic and predictive furnace model, with parameters determined experimentally at the site. The controller maximises the concentrate feed and keeps the furnace operational parameters inside limits given by raw material mix and equipment constraints. The products of a zinc roaster, calcine, sulphur dioxide and the side products, such as mercury, meet gradually more stringent requirements, in particular as to their assays and total recoveries in the process chain. The end users of sulphuric acid in northern Europe, among others in the pulp and paper industry, require mercury levels much less than the conventional 0.2-0.4 ppm Hg. Therefore, secondary de-mercurising circuits have been used over the last five years for further cleaning the SO2 stream allowing production acid grades with less than 0.08 ppm Hg.

Jan 1, 2008