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By Raymond Mancha

When calculating the pressure-volume characteristics of projected mine-ventilating circuits by orthodox methods, certain basic assumptions are required in order to employ the various available empirical data. It is assumed, for example, that the mine air is an incompressible fluid subjected to isothermal flow, an assumption sufficiently accurate for practical purposes since pressure and temperature differentials are small throughout the average circuit. A more erroneous assumption, however, is that all air is accounted for as it travels throughout the mine. No attempt is usually made to evaluate stopping leakage as it occurs; instead, the air volume required at the last crosscuts is assumed to enter the mine, travel the various intake air courses intact, sweep the workings and travel out by the return air courses to the point or points of exit from the mine. Actually, there is a leakage of air from intake to return at every stopping, the quantity of this leakage depending upon the tightness of the stopping and the pressure difference across the stopping. Unfortunately, a lack of empirical data and knowledge of the condition of individual stoppings makes an exact analysis of underground stopping leakage impossible. Generally, leakage is most severe through the old stoppings outbye the circuit. These are also subjected to higher pressure differences than the newer inbye stoppings. Therefore the circuit air volume diminishes at a decreasing rate progressing from outbye to inbye the circuit. Consider a simple mine circuit consisting of one intake and one return air course, each of equal and uniform section area and interconnected at regular intervals with crosscuts equipped with stoppings. At any point X inbye the circuit the air volume q on the intake air course equals that on the return air course and decreases with diminishing rate from outbye to inbye the circuit between the limits of Q1 completely outbye and Q completely inbye. Such a volume change is represented by curve a, Fig. I, which is computed with the aid of the equation q = Q1- (Q1- Q)(X + L)N. This equation results from the general algebraic equation of form q = k, — kXN after solving for kl and k with the knowledge that completely outbye X = o and q = QL, also completely inbye X = L and q = Q. Curves a, b and c, Fig. I, show values of q as ordinates versus values of X as abscissas, for values of N equal to 0.50, 1.00 and 2.00, respectively. Distribution of underground stopping 1eakage directly affects the air-volume distribution along the circuit, which in turn determines the circuit pressure gradient. To deliver any required air volume inbye a circuit or portion thereof requires a higher circuit-pressure differential if stoppings leak than is necessary with tight stoppings. Every mining engineer engaged in the analysis of projected mine-ventilation circuits is therefore confronted with the problem of making proper allowance for the effects of leaking stoppings. The purpose of this paper is to present evidence in support of the author's suggestion that circuit pressure increase (per cent)

Jan 1, 1942

By Raymond Mancha

When calculating the pressure-volume characteristics of projected mine-ventilating circuits by orthodox methods, certain basic assumptions are required in order to employ the various available empirical data. It is assumed, for example, that the mine air is an incompressible fluid subjected to isothermal flow, an assumption sufficiently accurate for practical purposes since pressure and temperature differentials are small throughout the average circuit. A more erroneous assumption, however, is that all air is accounted for as it travels throughout the mine. No attempt is usually made to evaluate stopping leakage as it occurs; instead, the air volume required at the last crosscuts is assumed to enter the mine, travel the various intake air courses intact, sweep the workings and travel out by the return air courses to the point or points of exit from the mine. Actually, there is a leakage of air from intake to return at every stopping, the quantity of this leakage depending upon the tightness of the stopping and the pressure difference across the stopping. Unfortunately, a lack of empirical data and knowledge of the condition of individual stoppings makes an exact analysis of underground stopping leakage impossible. Generally, leakage is most severe through the old stoppings outbye the circuit. These are also subjected to higher pressure differences than the newer inbye stoppings. Therefore the circuit air volume diminishes at a decreasing rate progressing from outbye to inbye the circuit. Consider a simple mine circuit consisting of one intake and one return air course, each of equal and uniform section area and interconnected at regular intervals with crosscuts equipped with stoppings. At any point X inbye the circuit the air volume q on the intake air course equals that on the return air course and decreases with diminishing rate from outbye to inbye the circuit between the limits of Q1 completely outbye and Q completely inbye. Such a volume change is represented by curve a, Fig. I, which is computed with the aid of the equation q = Q1- (Q1- Q)(X + L)N. This equation results from the general algebraic equation of form q = k, — kXN after solving for kl and k with the knowledge that completely outbye X = o and q = QL, also completely inbye X = L and q = Q. Curves a, b and c, Fig. I, show values of q as ordinates versus values of X as abscissas, for values of N equal to 0.50, 1.00 and 2.00, respectively. Distribution of underground stopping 1eakage directly affects the air-volume distribution along the circuit, which in turn determines the circuit pressure gradient. To deliver any required air volume inbye a circuit or portion thereof requires a higher circuit-pressure differential if stoppings leak than is necessary with tight stoppings. Every mining engineer engaged in the analysis of projected mine-ventilation circuits is therefore confronted with the problem of making proper allowance for the effects of leaking stoppings. The purpose of this paper is to present evidence in support of the author's suggestion that circuit pressure increase (per cent)

Jan 1, 1942

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'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 M. Prats

This paper presents a method for obtaining the correct values of the breakthrough sweep efficiency of the staggered line drive, taking into account the proper shape of the breakthrough streamline. Results obtained were different from those previously published, which were obtained by using an inexact shape of the breakrhrough streamline. INTRODUCTION In Section 9.31 of Muskat's' The Flow of Hotnogeticous Fluids Through Porous Media, it is stated that in a staggered line drive flood "the streamline of highest average velocity [is] the line of centers between input and output wells" (see Fig. I). Muskat used this concept to determine the relationship between the breakthrough sweep efficiency and the geometry of the staggered line drive pattern. Although it is clear from symmetry considerations that the breakthrough streamline passes through the midpoint between input and output wells, it is equally clear, as shown in Fig. 239 of Ref. 1, that the breakthrough streamline is not represented very well by the straight line joining the input and output wells. Therefore, it is desirable to determine the proper relationship between the breakthrough sweep efficiency and the pattern geometry by using the proper shape of the breakthrough streamline. METHOD OF SOLUTION It was found that the effect of the geometry of the staggered line drive on the breakthrough sweep efficiency can be investigated easily by making use of the method developed in Ref. 2. With this method, it can be shown that the general complex potential for the staggered line drive, which can be considered as a rectangular five-spot, is given in the coordinate system of Fig. 1, by O(z) = In cn(z\m) = +(x,y) + i $(x, y) . (1) where z = x + iy, +(x, y) is the potential distribution and $(x, y) is the stream function. The function cn (zm) is a Jacobian elliptic function of parameter m and has simple zeros at z,. = (2n + 1) K + 21iK and simple poles at z, = 2nK + (21 + 1)iK. Thus, the complex potential behaves like a(z) = ln(z - zr.) in the neighborhood of the production wells and behaves like a(z) = - ln(z - Z,) in the neighborhood of the injection wells. Since (z - z1.v) = rll.ei, we note that the potential + has the proper logarithmic behavior near the two wells, while the stream function $ varies with the angle, decreasing clockwise around the production well and increasing clockwise around the injection well. The functions K(m) and K'(tn), related by K(m) = K'(m,), where m + m, = 1, are known as the complete elliptic integrals of the first kind. By use of other properties of the Jacobian elliptic functions, the general stream function distribution for the staggered line drive can be obtained from Eq. 1: From the Jacobian elliptic function values, sn(0lm) = 0 dn(O(m) = cn(0jm) = 1, and f[K'(m)Iml = 03, it follows that the stream function given by Eq. 2 satisfies the proper symmetry conditions near a production well P located at [K(m), 01 and an injection well I located at [0, K' (m)]. That is, the line IOP in Fig. 1 is a streamline with value G (0. 0) = -0. and the straight lines connecting I and P to the point [K(m), K' (m)] is a streamline of value $[K(m), K' (m)] = -7/2. Similar symmetry conditions are obtained near other wells. It should be nuled that for m = m1, = 0.5, Eq. 2 reduces to the stream function distribution of the regular five-spot given in Ref. 7. SHAPE OF THE BREAKTHROUGH STREAMLINE It is apparent, from symmetry considerations, that the breakthrough streamline ($,) passes through the point [K(m)/2, K' (m)/2]. Its value there is so that the shape of the breakthrough streamline $,, is given by

Jan 1, 1957

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 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 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 R. D. Rein

The Henderson mine is a large continuous panel caving operation in central Colorado. The orebody is roughly 760 m by 910 m and will be extracted from two production levels. Each level is divided into three north - south panels 244 m to 317 m wide. The general arrangement of a production level and its associated undercut drifts is shown in Figure 1. The production drifts (3.6 m by 3.7 m) are spaced 24 m apart center to center and have undercut drifts (3.4 m by 3.8 m) 17 m directly above them. The production drifts are interconnected with drawpoint crosscuts (4 m by 4.3 m) driven on 12 m centers. Drawpoint crosscuts receive substantial brow steel and concrete support prior to undercutting. The brow steel is a 15 cm by 46 cm wide flange beam with a 2.5 cm wear plate to box in the 'cave' side of the beam. The concrete is a minimum of 0.3 m thick and the batch design results in 27 500 kPa strength. The finished drawpoint opening is 3.7 m by 3.2 m with a choke area of 3.7 m by 2.4 m. (Figure la) Undercutting and bell formation is accomplished by blasting rings of 7.6 cm holes drilled from the undercut drifts. The rings have a burden of 2.0 m and six rings develop one complete bell.

Jan 1, 1985

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 S. Nenno, T. Saburi, Y. Mizutani, M. Yomarnoto, K. Komatsu

In the equilibrium phase diagram of the Ni-Mo sys-tem1,2 available at present, intermetallic compounds (or ordered phases)ß(Ni4Mo) and ?(Ni3Mo) are known to exist to the nickel rich side. In our recent investigations on Ni-Mo alloys, however, it has been found that the quenched alloys from the high temperature region containing more molybdenum than the Ni4Mo do not decompose directly into 0 and when annealed isothermally below the peritectoid temperature, i.e., 865°C. This note is to report the existence of Pt2Mo type superlattice (possibly ordered Ni2Mo) as a metastable phase in the early stage of decomposition. Alloys containing 20.1, 22.7, and 24.4 at. pct Mo were prepared by vacuum-melting electrolytic nickel of 99.9 pct purity and molybdenum powder of 99.9 pct purity in an alumina crucible. From the resulting ingots, sheets 0.3 mm thick were obtained by forging, hot-, and cold-rolling. The sheet specimens were water-quenched from ! region and subsequently annealed for various periods of time at temperatures below the peritectoid temperature. On these specimens were made optical-microscopic and transmis-sion-electron-microscopic observations. Fig. 1- Optical micrograph of the 22.7 at. pct Mo specimen annealed for 100 hr at 700°C after quenching. (ß + ?) region, B, nucleated along a grain boundaries are invading the preexisting (ß + Ni2Mo) region, A. Details of A are observable only by electron microscopy. Etchant: 60 pct HNO3-40 pct CH3COOH.

Jan 1, 1970

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. 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 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 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'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 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 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 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

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 Jinwu Kang

The hydraulic turbine blade of the matensitic stainless steel (Cr13%,Ni5%,Mo1%) is a kind of large and curved shape, which is susceptible to deformation in both casting and heat treatment processes. The coupling analysis of heat transfer, phase transformation and stress in heat treatment process of the turbine blade for Three Gorges Project is carried out by using Deform-3D software. Its temperature fields, phase fields and stress fields in heat treatment process are obtained. And its deformation is predicted. In the cooling process of normalizing, the blade exhibits relatively large deformation. Based on the stress analysis and deformation prediction in heat treatment process, an algorithm for the determination of reverse deformation of the blade casting in heat treatment process is presented. And the reverse deformation distribution of blade is obtained.

Jan 1, 2010

According to the previous research, it is found that the thickness of cover is mostly less than 100 m for highwall mining system. However, for most of the Chinese surface coal mines, the mining depth is over 150 m; on the other hand, the shovel-truck is the most common mining technology. As a result, a lot of residual coal will be remained around end-walls. Considering these facts, the calculated methodology of highwall mining system under thick cover is presented in this paper. Meanwhile, numerical simulation of highwall mining system under thick cover was carried out. The outcome shows that the calculated methodology for pillar stability of highwall mining system under thick cover presented in this paper is acceptable. In other words, the highwall mining system is feasible to extract the residual coal around end-walls in Chinese surface coal mines. Thus, not only coal recovery ratio can be improved, but also can increase the production life of surface coal mines in China.

Jan 1, 2012