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By A. J. Smethurst, S. B. Hearn, S. Boskovic

"The problem of crud in solvent extraction is well known and can be a major handicap and operational issue. Colloidal silica can cause a number of problems including interfering with flocculation, initiating crud formation in solvent extraction (SX), decreasing SX extraction kinetics and increasing SX phase disengagement times. Flocculation can be utilized to remove some of the troublesome particulates but colloidal silica does not flocculate well and presents a different problem requiring innovative chemistry to solve. Coagulants can assist by facilitating the removal of this colloidal silica by promoting the agglomeration of the particles to a size which can be more readily settled or flocculated in solid/liquid separations. This is achieved when the coagulants bind to the surfaces, displacing water and destabilizing the colloids, whereupon they bind to each other which facilitate their amenability to conventional flocculation.INTRODUCTIONSilicate minerals are abundant in the upper layers of the Earth’s crust, and there are few mineral deposits without significant levels of gangue silicates. Silicate minerals include the pure silicas, such as quartz and other silicate containing minerals such as clays, chlorites, feldspars, as well a vast number of less common silicates.During the acid leaching of ores, two types of fine silica particles are released in solution from the dissolution of the host mineral. The first consists of partly depleted silicates, and the second is colloidal silica, which forms in solution. It is this second type which has the potential to cause crud formation as well as poor solution clarities, slow filter rates and excessive flocculant usage (Figure1)."

Jan 1, 2016

By David T. Peterson

Calcium metal was found to deoxidize thorizcm at 1000° to 1200° C. The reaction kinetics were determilled and related to the diffusion coefficients of oxygen in thorium. The solubility of oxygen in thorium, the minimum oxygen concentration, and the diffusion coefficient were determined from 1000° to 1200°C. This firocess results in the lowest oxygen concentrations zohich have been reported for thorium metal. FOR many years it has been known that calcium metal will reduce thorium oxide to thorium metal. This reaction has been the basis for several methods of preparing thorium metal. From the equations giv-by Kubaschewski and vans, ' AF" for the reaction Cao, + Tho,(,) - CaO(,, + Th(,) was calculated and found to be -3.4 kcal at 1000°C, -2.5 kcal at llOO°C, and -2.0 kcal at 1200°c. Thorium is very slightly soluble in liquid calcium, and the solubility of calcium in solid thorium is very low. Consequently these metals would be in essentially their reference states. If thorium containing oxygen were equilibrated with liquid calcium between 1000° and 1200°C, the oxygen content of the thorium would have to be below the solubility limit in thorium. Oxygen is one of the impurities most difficult to remove from thorium and is the most abundant impurity in metal prepared by almost all known methods. Fortunately, oxygen does not have a large influence on the properties of thorium because the solubility in solid thorium is very low. EvenT in thorium containing 100 ppm of 0, particles of thorium oxide can be observed in the microstructure. In view of the incompatibility of thorium oxide and liquid calcium and the low solubility of thorium oxide in thorium, the deoxidation of thorium by this method was investigated. For thorium containing an amount of oxygen well in excess of the solubility limit, the reaction should proceed in the following sequence. The oxygen content of the thorium matrix near the surface would be depleted by the diffusion of oxygen to the surface. At the surface, the oxygen would react with calcium to form calcium oxide. To maintain equilibrium within the thorium, thorium oxide would dissolve to keep the matrix saturated. Consequently, the thorium-oxide particles would disappear first at the surface and then the particle-free rim would grow in thickness. If the rate-controlling step were the diffusion of oxygen through this layer of thorium which was growing in thickness in direct proportion to the amount of oxygen removed, the well known parabolic time law should be observed. If the oxygen concentration at the surface of the thorium and at the inner surface of the deoxidized rim were known, the diffusion coefficient of oxygen in thorium could be calculated from the parabolic rate constant. EXPERIMENTAL PROCEDURE The thorium metal used in this study was prepared by calcium reduction of ThF, by the method described by Wilhelm.' The analysis of this metal is given in Table I. The carbon was determined by combustion, the oxygen by the HC1-insoluble residue method, nitrogen by the Kjeldahl method, and the other elements by emission spectroscopy. A section of this ingot was hot rolled at 600°C to 1/4 and 1/8-in. thick plates. Specimens approximately 7/8 in. square were cut from these plates, and all surfaces of the specimens were cleaned and smoothed by filing with a clean file. Individual specimens were placed in 1-in. diam by 2-in.-long tantalum capsules. Approximately 1 g of clean, high-purity calcium was placed in the capsules and an end closure arc-welded in place. The tantalum capsules were sealed in Inconel crucibles to protect the reactive metals from oxidation. The entire loading procedure was done in a glove box filled with pure argon. The loaded crucibles were placed in a muffle furnace, controlled within 2OC of the desired temperature, for a measured length of time. After the specimen had cooled to room temperature, it was sectioned perpendicular to the large faces and through the mid-point of two of the sides. The sectioned specimen was mounted and polished through Linde A abrasive. The rim which was free of thorium-oxide particles could be clearly observed microscopically as mechanically polished. Twenty measurements of the thickness of the rim were made at equally spaced points far enough from the end of the

Jan 1, 1962

By John R. Hutchins

1ms is a report of nickel (wire) oxidation studies designed to test the validity of Valensi's1 equation developed for the kinetics of a reaction controlled by diffusion in a continuously growing product layer on cylindrical surfaces. The equation is:

Jan 1, 1968

I AM sure you know that it is necessary to draw the gold into wire in order to make gold cloth, gold embroidery, or gold filigree work. Just as it is beaten to form leaf for ornamenting pictures, so also can gold be drawn easily. The same is done with silver and tin, and I believe it could also be done with iron, copper, and brass. Although brass is not so

Jan 1, 1942

By J. M. Font, R. G. Reddy

Jan 1, 2004

By John P. Shubilla1, Nicholas Kyriazi

A joint effort by the Pittsburgh Research Center (PRC) and the Mine Safety and Health Administration (MSHA)3was undertaken to determine how well self-contained self-rescuers (SCSR's), deployed in accordance with Federal regulations (30 CFR 75.1714), held up in the underground environment with regard to both physical damage and aging. This report presents findings regarding laboratory-tested SCSR's in the fifth phase of testing from mid-1993to early 1996. The SCSR's were tested on human subjects and on a breathing and metabolic simulator (BMS) at PRC. These results indicate that most of the apparatus, if they pass their inspection criteria, perform satisfactorily.

Jan 4, 1996

By Earl L. H. Sackett

Barite, although not a glamour mineral and probably little known to many of those in the mining business, is produced in the US. in very respectable quantities and is an important factor in the mineral economy of this country. The scope of the various uses of this mineral is perhaps one of the most misunderstood things about it, for it is surprising how many of those with a casual knowledge of the subject speak of its primary utilization in terms of paint pigments and barium chemicals. The life blood of barite utilization, however, is the petroleum industry by virtue of its dependence on this high specific gravity, inert mineral as the weighting material in its drilling fluids. From 80 to 85% of barite usage in the U. S. is accounted for in this manner-in fact, its role in oil well drilling is the only excuse for some of the large-scale barite production facilities which have been developed. Other uses are important and widely varied, but comparatively they are of small scale. They include fine crushed barite of low iron content as a component in glass manufacture, barite as a source material for manufacture of barium chemicals, finely divided barite as a component of some rubbers and ground, bleached barite in pigments. The formerly substantial use of barite in lithopone, however, has declined markedly because of replacement of lithopone by titanium pigments

Jan 5, 1962

By M. Yao, D. Thibodeau

"Selective mining, such as post pillar cut and fill, is used in large, flat, irregularly shaped ore bodies where ground stability is to be maintained. This selective mining is used to open the stope to the full lateral extent of the ore body in order to minimize the dilution and maximise the recovery. For increased mining depth, design of the post pillar becomes more challenging since those pillars will eventually fail as mining progresses upwards. The post pillar should be designed such that failure does not occur at the early mining stage, such as at the first cut. After mining progresses upwards, the post pillars are expected to fail but be confined by backfill. In order to maximise ore recovery, both drift width and post pillar size must be designed adequately with a consideration of all geomechanic parameters. This paper will describe the process used to design post-pillar cut and fill for one of the ore bodies in Vale’s Copper Cliff Mine using a combination of both empirical deterministic and numerical modelling techniques. By utilizing both empirical deterministic and numerical modelling techniques drift width and post pillar size can be effectively designed. Current post pillar cut and fill design in the existing Vale operations are used to validate the design for Copper Cliff Mine. INTRODUCTIONCopper Cliff Mine is located within the Copper Cliff Offset in the limits of City of Greater Sudbury, Ontario, Canada, as shown in Figure 1. The Copper Cliff Offset extends about eight kilometres south from the Sudbury Igneous Complex into the footwall rocks. Until 2008, the current Copper Cliff Mine was operated as two separate mines i.e. North Mine and South Mine. However, as part of restructuring of Vale Ontario Operations, the North and South Mines were merged in 2009 to form the “Copper Cliff Mine”."

Jan 1, 2015

By Marc Symbaluk

Core Programs Current ? Aboriginal Involvement Program ? Adaptive Forest Management / Forest History Program ? Grizzly Bear Research Program ? Natural Disturbance Program ? Foothills Growth and Yield Association ? Foothills Stream Crossing Program ? Fish and Watershed Program ? Caribou Landscape Management Association ? Social Science Program ? Local Level Indicators Program ? Communications and Extension Program ? Geographic Information Systems Program ? Mountain Pine Beetle Ecology Program In Development ? Climate Change Program ? Forest Community Program ? Water, Quantity and Quality Program (Water for Life) ? Circumboreal Initiative ? Yellowhead Ecosystem Management Group

May 1, 2008

By Elton A. Youngberg

The Holden mine operated by the Chelan Division of the Howe Sound Co. is on the east slope of the Cascade Range in north central Washington on the south slope of Railroad Creek valley at an elevation of 3500 ft. The mine may be reached by a 40 mile boat trip from the town of Chelan which is at the southern tip of Lake Chelan, to Lucerne at the mouth of Railroad Creek and an 11. mile bus ride up Railroad Creek to Holden. A11 freight and concentrate is moved over this route to Chelan Falls on the Columbia River which is on the railroad four miles below the town of Chelan. The mine is now producing 2000 tons of gold, copper, and zinc ore per day which is treated in the Holden mill. Gold-copper and zinc concentrates are made, the first of which is shipped to Tacoma, Wash., and the latter to Kellogg, Idaho, for smelting. Ore is broken by long-hole blasting using the Noranda system which has been modified to meet local conditions. Until recently, blastholes have been drilled by diamond drills. Now a partial substitution of percussion drill holes, drilled with tungsten carbide insert bits, is being made. Geology The ore body occurs as a replacement deposit in a highly metamorphosed series of sedimentary rocks, mainly gneiss and schists, in a shear zone several hundred feet in width and of undetermined length. Commercial ore has been found in mineable widths of 25 to 100 ft for approximately 2500 ft along its strike. The commercial minerals are chalcopyrite, sphalerite, and gold. During the period of mineralization considerable silicifica-tion took place giving the ore an abrasive drilling characteristic. Following the period of mineralization, numerous dikes were introduced into the ore body. The earlier ones were of granite composition having a width of a few inches up to 80 ft. These were followed by much younger, fine grained basic dikes which usually do not exceed 2 ft in width. Development of Percussion Blasthole Drilling Equipment Test work with the 1½-in. tungsten carbide bit was carried on in development headings for several months early in 1947. The short life of the bits, because of gauge loss caused by the abrasive nature of the rock, prevented its adoption for this use. However fast drilling speed and ability to drill a long uniform hole suggested its use for drilling blastholes in competition with diamond drills as diamond costs were steadily increasing and exper-ienced drillers were difficult to obtain. The 1½-in. bit was the largest available at the time initial test work was started with sectional steel. The 1½-in. hole limited the diameter of the steel thread and coupling which could be used. Type F couplings were first used but because of the small thread section excessive breakage of the steel was experienced. Type H couplings were tried next. In order to use this coupling which is 15/8 in. in outside diameter, it was reduced to 1 3/8 in. giving 1/8 in. clearance between the coupling and the hole. Rod breakage at the thread was substantially reduced but some coupling breakage was experienced, however the overall performance was considered satisfactory (see Fig 1 for illustration of coupling and thread). Early test work with the 1½ in. bit indicated machines of piston diameters larger than 255 in. would cause inserts to loosen or break. It was found however that the additional weight of the sectional steel cushioned the blow enough to prevent bit failures when 3-in. Leyners were used. Rods used with the 1½-in. bits were 7/8 in. q. o. for sectional steel and 1 in. q. o. for all chuck pieces. In May 1948, 2-in. tungsten carbide bits became available and test work was immediately started. The 2-in. hole approximated the AX (1 15/16 in.) diamond drill hole which was being used exclusively for blastholes and permitted their substitution for diamond drill holes in a ring without alteration of pattern, burden, or explosives. The 2-in. bit also gave room in the hole for larger couplings and permitted the use of heavier rods and 3½-in. machines, increasing the

Jan 1, 1950

By Wern-ping Chen

Technical and non-technical challenges exist in each mega infrastructure project. It?s no exception to the first US underground PPP Port of Miami Tunnel (POMT) project in Miami, FL. The POMT will connect two man-made islands, the Dodge and the Watson Islands, under the Biscayne Bay within highly porous limestone formation. Permeability as high as 0.1 cm/sec has been identified. Its 41-ft outside diameter, if constructed, is the largest precast segment lining in the US. Besides its geological and hydro geological challenges and risks, local community influence and impacts have controlled the fate of this project. This paper describes its tender design enhancements and the evolving of its slow advancement in resolving local community issues.

By N. S. Yet

This paper shall present the design and construction of bored pile and diaphragm wall for the proposed MHA complex in Singapore. The work was carried out by L&M Foundation Specialist Pte Ltd and PWD Consultants Pte Ltd was the consultant for this project. A brief description of the project is first presented to illustrate the scope and scale of the work involved, which includes the construction of bored piles and diaphragm walls and installation of steel stanchions in the bored piles. This is followed by a description of the methodology used to analyse and design the wall and bored piles. As both towers in this project are supported on bored piled foundation and are constructed by top down construction method, the analysis of the wall has taken into consideration the sequence of work phase by phase and this was carried out by means of a commercial finite element program called CRISP. Modelling of the problem will be briefly described and the findings from the numerical analysis will be presented. This is followed by a description of the structural and geotechnical design aspects of bored piles. The paper shall then describe the difficulties encountered for this project and solution used to solve the problem. The subsoil condition will be described with special mention on the encounter of very strong granitic rock at depth above the maximum excavation level of the basement at certain locations, hence necessitating the drilling of rock for a considerable depth for both bored piles and diaphragm wall. At this juncture, the construction method used to drill through the rock will be briefly described. For construction of bored piles, penetration of rock was carried out by chiselling or drilling using reverse circulation rigs. For construction of diaphragm wall, it requires the employment of Bauer hard rock cutter which uses cutting wheels fitted with roller bits and are capable of excavating into granitic rock of compressive strengths more than 100 MPa.

Jan 1, 2000

By William H. Kobbé

The word recovery as used in this paper is applied in its broader sense and not limited to wells producing from horizons of unconsolidated sands. Certain problems connected with the winning of petroleum from such horizons present themselves while drilling is in progress and should be considered before drilling is commenced wherever unconsolidated sands are expected. Maximum recovery depends not alone. upon efficient pumping methods but also upon the selection of a proper drilling system and the completion of a well especially adapted to the extraction of these oil-bearing sands. 11. DRILLING Determining Factors in Selection of Method The selection of a drilling method for the development of a region in which the oil reservoir consists of unconsolidated sands is based on the following considerations: (1) Character of the formations overlying the reservoir; (2) thickness of the oil-bearing stratum or strata; (3) thickness and character of water or gas sands; (4) gas pressure to be encountered or expected; (5) total depth of well. 1. Character of the Formations Overlying the Reservoir The strata overlying the reservoir may consist of beds of shales, clays, sands and other soft materials, or they may be slates, limestones and sandstones consolidated to all degrees of hardness. Very often the overlying strata are predominantly of soft materials but interbedded with harder formations. Likewise, soft strata are encountered at different horizons in an overburden consisting almost entirely of the harder sedimentaries. The physical character of this overburden is frequently

Jan 1, 1917

By B. Cavender

Mining managers trying to rapidly reduce production costs tend to implement a traditional set of top-down managerial practices: layoffs of employees, renegotiation of supplier contracts, cancellation of employee training, restrictions on travel, reductions in office supplies, and so forth. While these practices do reduce costs, the benefit is often limited in both magnitude and duration, and the workforce is usually left demoralized. Successful cost-cutters take a different approach. They make the continuous reduction of costs a priority for all members of the organization, whether managers or workers. The culture of these firms emphasizes the commitment, voluntarily given, of individuals and work groups to this goal. Cost cutting practices often reject all of the traditional methods mentioned above, but are far more numerous, innovative, and successful. This culture of cost reduction is the result of a set of leadership competencies that establish direction for the organization, align the organization’s various constituencies with this direction, and motivate and inspire the organization. Dramatic cost reduction then becomes a routine task rather than a heroic achievement.

Jan 1, 2000

By Wilbur I. Duvall

The borehole deformation gage and overcoring techniques developed by the Bureau of Mines form one of the primary methods of determining in situ rock stress. This report describes a least squares method of calculating the average rock stress components in the horizontal plane from more than three diametral deformation measurements in a vertical borehole. This least squares analysis provides a means of determining the multiple correlation coefficient, the variance, and the scatter for a set of deformation measurements. From this information, confidence limits can be calculated for the secondary horizontal principal stresses and their directions. The material presented in this report can be used as a guide to the development of least squares solutions for stress measurement techniques other than the borehole deformation method. Since the statistical information that results from the least squares method is extremely useful in the interpretation of results, it is recommended that a least squares analysis be used on stress measurement data whenever possible.

Jan 1, 1979

By William A. Beck

"The Piedmont Manganese Belt extends from Taliferro County, Georgia, 70 miles in a northeasterly direction through Greenwood County, South Carolina, and conforms with the trend of the Piedmont Province. The total area of the belt is reported to be about 1,000 square miles, and the maximum-width is about 24 miles.Manganese has been known to exist in the belt since 1845. M. Tuomey mentioned a deposit in the Edgefield district on Hard Labor Creek in South Carolina in a report on the geology of South Carolina published in 1848.Ore was first mined about the time of the first World War, when the Colley mine, near Lincolnton, Ga., was. Opened. The only operation in the area at present is that of the Abernathy & Kehaya Mining Co., at McCormick, S. C.The geographical position of the belt is from west longitude 82° 57’, north latitude 33° 39' to west longitude 82° 06', north latitude 34° 08'. The towns of Washington and Lincolnton, Ga., and McCormick and Greenwood, S. C., are in or near the belt.An exploratory project of the Bureau of Mines was started on the McCormick, S. C., deposit in April 1944, and trenching, test pitting, and hand-auger drilling-were done on a small scale until early in October 1944. (Fig. 1.) In November 1944 work was begun on a property 12 miles north of McCormick, known as Parsons Mountain, which is leased to M. H. McKnight by the United States Forest Service. From March 15 to March 24, 1945, work was done on the property of Mrs. H. B. Ross, which is known as the Morgan farm.The Abernathy & Kehaya manganese mine contains what is locally known as the Dorn gold mine; mineral rights on the property were owned by M. G. & J. J. Dorn, Inc."

Apr 1, 1946

By Laurence D. Burke

The electrocatalytic behaviour of gold in aqueous media is surprisingly complex and a novel view of gold electrode surfaces is proposed to explain the observed electrochemical responses. The metal surface is considered as a chemically (or redox) modified electrode, with equilibrated, low energy, surface gold atoms functioning as a relatively inert support and low coverage, high energy, protruding gold atoms (or minute clusters of same) functioning as electrocatalytic redox mediators. The protruding atoms are viewed as surface active sites and the complex electrocatalytic behaviour is attributed to the basic instability of the mediators, the diversity (and poor control) of same, and the acid/base (or hydrolysis) behaviour of the oxidized states of the mediators. It is assumed that major advances in the application of gold in the heterogeneous catalysis of gas phase reactions will be accompanied by similar advances in the use of gold in the electrocatalysis area; however, this will require considerable work of both a practical and theoretical nature.

Oct 1, 2003

By R. A. Sharma, F. D. Richardson

Manganese oxide acitivities and sulfide capacities have been measured at 1650°C in mixtures of MnO + A1203 and MnO + SiO2 - Al2O3. Sulfide capacities have also been measured at 165O0C for melts of MnO + SiO2, MgO + SiO2, MgO + SiO2 + Al2O3, MgO + MnO + SiO2, and SiO2 + Al2O3 and for crystalline CaO. The activity coefficients of the sulfides of manganese and magnesium have been derived where possible. It looks as though the activity coefficients of sulfides are greatest at compositions where the free energies of formation of the silicate mixtures have their maximum negative values. The amphoteric nature of alumina is reflected in the results for the binaries and ternaries in which it is present with silica. The limiting solubility of CaS in CaO at 1650°C is 6.3 x 10-4 mole fraction. THE activities of manganese oxide have been measured in a variety of silicate and aluminate melts by use of the equilibrium (MnO)slag = [Mn]pt .½O2 [l] and the experimental techniques described previously.' Sulfide capacities (CS = wt pct S (PO2/PS2)½) have also been determined in a series of melts, by use of an established method2 which exploits the equilibrium (O°) + ½S2 = (S") . ½O2 [2] If CS and the activity of a metal oxide are known in a melt, the activity coefficient of the metal sulfide can be calculated by use of the equilibrium constant K, for the pure phase reaction MO i ½S2 = MS + ½O2 [3] The activity coefficient of the sulfide is then given by the equation

Jan 1, 1965

By Jochen Petersen, Harro von Blottnitz

"The landfill disposal of slags is common practice in the metallurgical industry. Chromium may be a minor constituent in many cases. It is present mostly in the form of chromites, which are considered environmentally benign and hence safe for disposal. Wastewaters from metal finishing operations are most commonly treated by reduction and precipitation of dissolved metals, including chromium, by lime addition, and subsequent landfill disposal of the resulting sludge. The metals contained in such sludges are mostly in the form of hydroxides, and as such considered environmentally stable.Thermodynamically, trivalent chromium, which is the predominant form in these waste materials, can be oxidised to the hexavalent form by atmospheric oxygen. Detailed research work on both slags and sludges has shown that a gradual emergence of hexavalent chromium can indeed be observed. The mechanism is governed by close physical contact between chromium oxide (or hydroxide) phases on the one side and calcium oxide (or hydroxide), which is present as flux in slags and residual lime in sludges, on the other. If oxygen can migrate to the points of contact, gradual formation of calcium chromate ensues.Hexavalent chromium is associated with high environmental toxicity, and hence undesirable in landfilled materials. Moisture migrating through landfill deposits can mobilise the highly soluble chromates and thus potentially transport them into natural waterways. Such gradual release from stainless steel and ferrochromium slag deposits has been demonstrated in a column experiment."

Jan 1, 2003

By Chris Searing

An underground room-and-pillar limestone mine in Western Missouri acquired a new primary crusher, physically larger in every dimension than the unit it replaced. Due to the space constraints within the mine, resetting the truck dump and feeder to account for the larger crusher was unfeasible. Mine management determined the most practical solution was to increase the available area for the new crusher via blasting. Fitting the new crusher into the existing area required the removal of 6 feet (1.83m) of limestone from the original floor grade. Blast parameters had to be designed to effectively break this shallow depth of rock while controlling movement to avoid contacting the mine back 23 feet (7.01m) overhead, an apron feeder 13 feet (3.96m) above the southern extent of the area, and the original crusher which was still in place 15 feet (4.57m) from the first pattern. Additionally, work needed to be completed within an aggressive schedule set forth by the mine, allowing little room for delay or error. This paper will examine the efforts taken to safely and effectively blast the rock while protecting nearby equipment and integrity of the mine back. Subjects covered will include blast design, adjustments made to account for drilling and equipment-related constraints, and how seemingly small changes in load parameters can significantly impact performance with small-diameter blast design.

Feb 6, 2023