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By R. P. Schneider

The Chino mine, concentrator, and smelter are located in Grant County in Southwestern New Mexico. Chino Mines Company is a partnership two-thirds owned and operated by Kennecott Santa Fe Corp., and one-third owned by MC Minerals Corp. (Mitsubishi) of Japan. The Chino Ore Body The Chino ore body is a widely varied deposit of rock types and contains intrusive stocks and sediments. The stocks contain chalcocite in a matrix of quartz, mica, and feldspar. Much of the feldspar is altered to clay. Some of these stocks are enriched with native copper. The sediments contain chalcopyrite associated with gangue minerals including magnetite, pyrite, garnet chlorite, and epidote. The contact of the stock and sediment is heavily altered to chlorite and clay. The stock ores contain economic quantities of molybdenite. The operating work index of Chino ore varies from 5 to 20 kwh/ton and averages 8.5 kwh/ton. The ore head to the concentrator averaged 0.81% copper and 0.010% molybdenum. Fifteen percent of the total copper head exists as nonsulfide mineralization.

Jan 1, 1985

By P. K. Lucke

One of the interesting features connected with the great continental uplift, which formed the table land of Mexico, is the great depth to which oxidation and secondary enrichment of orebodies occurred. This table land lies at altitudes ranging from 4000 to 7000 ft. (1219 to 2133 m.) above sea level. Several cases could be cited of oxidized zones in copper and lead mines which reached depths of upward of 1500 ft. (457 m.) below the surface, and it is universally admitted that the oxidized zone in a mine is limited by the permanent water level. Mexico, however, like many other countries in which deserts occur, is visited at rare intervals by tremendous floods, which sometimes cause variable water levels over long periods. The following cases bear out the idea, and lead to interesting conclusions of commercial value. My paper demonstrates that carbonates of ores may occur below water level and sulfides above, and I will endeavor to explain the causes of the phenomena in certain specific instances. Case I.—Fig. 1 in this case shows that sulfides of lead remained

Jan 1, 1920

By Güven Önal, Zeki Dogan, Neset Acarkan

In this study, it was aimed to produce glass sand with less than 0.15% Fe2O3 for window glass manufacture and glass ware grade with less than 0.05% Fe2O3 from Yalõköy sand dunes having 0.48% Fe2O3. In order to reduce the iron content of the Yalõköy sand dune, magnetic separation, flotation, scrubbing, H2SO4 leaching and drying were used. The optimum iron removal was achieved by H2SO4 baking at 150oC and 8 kg/t acid consumption. In this way, sand concentrate was obtained with 0.04% Fe2O3 and sand recovery 95%.

Jan 1, 2003

By C. Chiou, D. P. Seraphim

The effect of clustering on the superconducting properties of Al-Zn alloys has been studied by the ballistic induction techniques. The superconducting critical temperature, T,, changes when zinc-rich clusters are formed. The change in T, depends on the concentration and may be positive or negative, but always correlates with an effective decrease in the solute concentration of the alloy. Reversion after clustering effectively returns a 5 at. pct Zn alloy to the quenched state. These methods might be use&lly extended to the Al-Mg system, where X-ray data are difficult to obtain. In very dilute Al-Zn alloys, as in other tin,' indium, aluminum,2 or tantalums base solid solutions, there is a decrease in the superconducting critical temperature, Tc, roughly proportional to the temperature independent resistivity, i.e., -0.2°K per pS2cm. With increasing solute concentration, the T, generally reaches a minimum and then increases4 to a value well above that for the pure solvent. The change of T, with solute concentration follows an empirical formulation involving two constants which correlate with the scattering cross section of the impurities. As will be shown below, the Al-Zn alloys more concentrated than 1or2at. pct do not follow the empiricism, probably because they are unstable even when rapidly quenched.5 However, it seems possible to study at least the later stages of solute clustering by precise measurements of the 1', following various thermal treatments. A few experiments on quenched alloys, aged alloys, and tests on reversion are reported here. The heat treatments selected for obtaining the quenched and aged states were mainly deduced from the work of Dash and Fine.6 For the quenched state, specimens were homogenized at 480°C, quenched to -160°C, and immediately transferred and mounted in liquid nitrogen. Measurements on the superconducting properties and residual resistivities were then made at helium temperatures, after which the specimens were aged at room temperature for 3 to 5 hr, i.e., much longer than the time required for clustering to be essentially completed. The superconducting properties and residual resistivities were once again examined. To reach the reverted state of the A1-5at. pct Zn alloy, the aged specimen was heated for 5 min at 200°C and then tested as above.

Jan 1, 1963

By Ian E. Efford

The paper gives an assessment of Canada's energy supply-and-demand position, the current situation and projections into the future, as well as a review of some alternative energy futures for Canada. The main elements of Canada's energy strategy for the period 1977-1990, with particular reference to the role of energy conservation, are discussed, leading in to the development of the national energy conservation program.

Jan 1, 1977

By A. Warczokand

The cleaning of copper smelting slag requires magnetite and cuprous oxide reduction. The slag reduction with coke, injected coal powder or bunker oil is strongly endothermic and requires heat generation by fuel combustion or the use of electric energy. The reduction of fayalitic slag with calcium carbide is slightly exothermic and the formed calcium silicate modifies the slag. Thus, the additional energy is required only for covering furnace heat losses. Results of crucible scale tests showed the rate of slag reduction approximately five times higher in comparison of the slag reduction with coke under the same conditions. It seems that the effect of exothermic reaction and slag modification with CaO at the reaction interface results in very high rate of slag reduction. Industrial scale tests carried out in the Anglo-American Chagres Smelter demonstrated the possibilities of effective reduction and cleaning of the slags from a flash smelter and Peirce-Smith converters in a Teniente Slag Cleaning Furnace. The slag reduction was carried out without any reductant or stirring gas injection, simply by calcium carbide charging through a furnace mouth, and was very fast.

Jan 1, 2007

By Thomas L. Kesler

IF all of the 14 million tons of kaolin produced in Georgia through 1949 had been mined from a single deposit 20 ft thick, it would represent a mined-out area of less than 1 sq mile. This measure of depletion is in such strong contrast with the large size of the area shown in Fig. 1, from which most of the output has come, that there has developed an optimistic outlook concerning future supplies of kaolin equal in grade to that already mined. Dozens of mines are widely but not evenly distributed within this area, which consists mostly of parts of Twiggs, Wilkinson, and Washington counties. There are scarcely half a dozen kaolin mines elsewhere in the state that have been appreciably productive, although the areas in which they occur may have good possibilities. Most of the efforts to find new deposits, therefore, will be confined to the area shown in Fig. 1, for it is here that the industry has established its facilities, and prospecting is encouraged by relatively numerous outcrops and opened deposits of kaolin.

Jan 10, 1951

By M. Clotilde Apua, A. F. Mulaba-Bafubiandi

"The efficiency of hydrochloric acid (HCl) leaching of gold from low grade gold-bearing material has been investigated in the presence of sodium hypochlorite (NaOCl) to produce in situ chlorine gas which is an oxidizing agent able to oxidize gold metal to its soluble forms. The effect of concentrations of HCl, NaOCl, and mixtures of HCl + NaOCl was sought. Iron and potassium present in the feed were leached and consumed all the HCl. Gold found as inclusion in pyrite was not exposed to the lixiviant. Therefore, leaching of gold from low grade gold-bearing material in aqueous chlorine solution is not an effective process for now.INTRODUCTIONHydrometallurgy of gold is commonly based on cyanide leaching solution. Due to the environmental concerns related to cyanidation and mercury processes, and inefficiency of cyanide gold leaching to dissolve refractory ores, sustainable alternative lixiviant systems, capable of generating gold (I) and gold (III) complexes must be studied. Amongst non-cyanide gold lixiviants there is the chloride/chlorine system (Hilson and Monhemius, 2006; Senanayake, 2004; Pilsniak and Trochimczuck, 2007; Wang et al., 2007; Reeves et al., 1999; Khodadad et al., 2008).In practice, the most oxidizing agents utilized for the leaching of gold as chloride, are chlorine (Cl2) or hypochlorite (HOCl) which are either generated in situ by suitable reaction or added in solution (Adamson, 1972). The dissolution of gold concentrate using hydrochloric acid, sodium chloride and sodium hypochlorite (Cl-/HOCl system) were studied (Puvada and Murthy, 2006). The effect of NaOCl and NaCl concentrations were also investigated. The leaching was performed under the following conditions; time: 4h, solid–liquid ratio = 1:4, at room temperature. It was found that the increase of the NaOCl concentration from 25 to 75 g/L did not have any meaningful effect on the recovery of gold which was about 30% (experimental conditions; S/L ratio = 1:4, NaCl concentration: 25 g/L, HCl concentration: 0.35 M). It was also found that from 25 g/L to 200 g/L NaCl, the recovery of gold rose from 30% to 40% (experimental conditions; S/L ratio: 1:4, NaOCl concentration: 25 g/L, HCl concentration: 0.35 M)."

Jan 1, 2012

By C. B. Post, G. V. Luerssen

J. STEVENS*—The authors state that the contamination in the metal was due mostly to the ladle refractories. Did the contamination vary with different brands of refractories or different qualities in the ladle ? C. B. POST (authors' reply)—That would be expected. In our shop we have used a grade of firebrick which is fairly standard. We do use high refractory nozzles to combat this erosion and high refractory sleeves, but there is such a problem when you get a skull in a ladle you just cannot use these high refractory bricks; you rip out the whole ladle getting the skull cleaned up. The acid bricks form a glass, and you can pull the skull out when you do get them. I believe you might find variations among the different refractories depending on how much silicon is in the brick. This determines how much silica is going into the interface between the steel and the brick. Our brick is about 35 pet alumina and 65 pet silica. With a brick of 35 pet alumina in the slag patches new calculations would have to be made. R. PARDEE*—Where can you draw the line in determining the refractories as being a silica refractory ? C. B. POST—More silica than about 50 pet. I think perhaps that is generally the criterion. The brick we use is a good high grade fire brick. T. S. WASHBURN*--With reference to the higher silicon that is recommended as a result of this investigation, is it contemplated that the specified range would be raised to 0.40-0.60 pet silicon, or possibly 0.50-1.00 pet silicon? These higher silicon ranges would affect the physical properties of the steel and consequently there might be some complications with respect to educating the consumers to 'accept this revision of the conventional analysis. C. B. POST—It will take considerable work. Take AMS 6260 as an example. Ordinarily this specification calls for 60 or 75 manganese. Silicon has some

Jan 1, 1950

By M Olsson, U Nyberg, K Widenberg, P Svedensten, F Ouchterlony

During 2007 and 2008, the project Optimal Fragmentation in Blasting II was conducted at the Långåsen quarry of NCC Roads AB near Arlanda. The main goals were:an evaluation of the effect of using electronic delay detonators (EDDs) in quarry blastinga validation of the design curve concept introduced at the Vändle quarry.Five production blasts were monitored. After a first trial round, the remaining four rounds each had about 100 Ø89 mm holes in a 14–19 m high bench charged with site mixed emulsions. In rounds 1 and 2, a tighter pattern than normal was used in one half of the rounds, raising the specific charge there from 0.7 to 1.0 kg/m3. Non-electric initiation was used with two holes per 25 ms delay in-row and 67 ms between rows. In rounds 3 and 4, electronic delay detonators were used instead, with 10 and 5 ms in-row delays respectively.The monitoring encompassed drill hole collaring and alignment, in-hole deviations, the rock structure in the bench faces from 3D imaging, first row burdens, etc. Eleven 400 t samples were taken from the muck piles and sorted in the quarry and sieved in the laboratory (lab). The lab data made up a fines tail that was grafted onto the in-pit sorting data to form sieving curves for the test piles. These curves were then compensated for the absence of boulders to construct the sieving curves for the blasted muck pile.These have been used to derive fragmentation model formulas for the non-electric rounds, also called design formulas. They are based on blast geometry, charging data and the properties of the rock mass. The formulas may be used to estimate the effect of changing blasthole pattern, borehole diameter and stemming length for example.The use of EDDs gave several unexpected results. In comparison with non-electric initiated rounds with a somewhat lower specific charge the fragmentation was much coarser; x50 increased from 160 to 200 mm. The effect of raising the specific charge in the non-electric initiated rounds from 0.72 to 0.99 kg/m3 was a much finer fragmentation; x50 decreased from 160 to 120 mm.CITATION:Ouchterlony, F, Nyberg, U, Olsson, M, Widenberg, K and Svedensten, P, 2015. Effects of specific charge and electronic delay detonators on fragmentation in an aggregate quarry – building fragmentation model design curves, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 727–740 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By Frank E. Chase

Kettlebottoms are columnar masses of rock--the preserved casts of ancient tree stumps--embedded in coal mine roof strata (of which they are a part). Because unsupported kettlebottoms can detach from a mine roof without warning, they are a hazard to miners. The primary objectives of this investigation were (1) to better define and describe kettlebottoms and (2) to evaluate kettlebottom support techniques in terms of safety and effectiveness. A telephone survey of mine operators, interviews with Mine Safety and Health Administration (MSHA) personnel, and underground mine visits were conducted in Pennsylvania, West Virginia, and Kentucky. Information gathered during this investigation indicates that the size and frequency of kettlebottoms in mine roof are dependent upon past geologic events and biological processes active during the deposition of roof sediments. To ensure the safety of mine personnel, all undermined kettlebottoms should be supported. The roof should be bolted next to kettlebottoms less than 3 ft in diameter, close enough to allow a portion of a wood or steel header to be extended beneath each kettlebottom for support. Two bolts and a wood plank or steel strap should be employed to support kettlebottoms over 3 ft in diameter.

Jan 1, 1983

By H. A. Doerner, F. D. DeVaney, J.B. Clemmer

Production of magnesium metal in the United States during the past decade has increased from less than 600,000 lb. in 1928 to more than 4,800,000 lb. in 1938.1 The growing industry has stimulated interest in methods of production and in sources of raw material. All of the present domestic production is from brines. However, the availability of large tonnages of magnesite in the state of Washington, together with the possibility of cheap power, have created interest in the utilization of magnesite for the production of magnesium metal as well as for other purposes. Magnesite, the carbonate of magnesium (MgCO3), usually is associated with limestone (CaC03) and dolomite (MgCO3. CaC03). It occurs in crystalline and microcrystalline (sometimes termed amorphous) forms.2 The more common crystalline mineral is found in various shades of white, gray, yellow and brown, the colors being due to impurities, chiefly iron. Magnesite has a hardness of 3.5 to 4.5 and a specific gravity of 3.0 to 3.1. It forms rhombohedra1 crystals in the hexagonal system, having a vitreous, pearly luster. The quarries and mines of the Northwest Magnesite Co. near Chewelah, Wash., are large producers of magnesite for refractories. The deposits are extensive, some veins being hundreds of feet wide. The boundaries of the veins are indefinite, and the magnesite is interfingered with dolomite. Calcite, talc, serpentine, quartz and shale are disseminated through the deposits. The quality of the ore cannot be determined by appearance or by simple tests; chemical analysis is required. These features make it difficult to determine the boundaries or total quantity of workable ore.3 Only the higher-grade ores are mined, and a considerable proportion of the magnesite content is discarded by hand-sorting and log-washing methods. After calcining, the sorted product meets the specifications for usual refractory material, as it contains approximately 81 per cent MgO, 7 per cent SiO2, 6 per cent CaO and 6 per cent Fe203 + A1203. A sample of the reject material had the following analysis after calcining: 65.4 per cent MgO, 6.3 CaO, 21.1 SiO2 and 6.5 R2O3. The percentage reported as R2O3 includes compounds whose chlorides are acid-soluble but give precipitates in ammoniacal solution that are ignited and weighed. The more common constituents are Fe2O3, A1203 and TiO2. The sorted magnesite product is not the best grade. even for refractory purposes, and leaves much to be desired for a reduction process. The rejected material contains a large proportion of magnesite. It is apparent that an effective concentration method would yield a higher recovery and a better product. In general, the refractory grade is not suitably pure for other uses. Recent developments in the production of

Jan 1, 1943

By J. B. Clemmer, F. D. DeVaney, H. A. Doerner

Production of magnesium metal in the United States during the past decade has increased from less than 600,000 lb. in 1928 to more than 4,800,000 lb. in 1938.1 The growing industry has stimulated interest in methods of production and in sources of raw material. All of the present domestic production is from brines. However, the availability of large tonnages of magnesite in the state of Washington, together with the possibility of cheap power, have created interest in the utilization of magnesite for the production of magnesium metal as well as for other purposes. Magnesite, the carbonate of magnesium (MgCO3), usually is associated with limestone (CaC03) and dolomite (MgCO3. CaC03). It occurs in crystalline and microcrystalline (sometimes termed amorphous) forms.2 The more common crystalline mineral is found in various shades of white, gray, yellow and brown, the colors being due to impurities, chiefly iron. Magnesite has a hardness of 3.5 to 4.5 and a specific gravity of 3.0 to 3.1. It forms rhombohedra1 crystals in the hexagonal system, having a vitreous, pearly luster. The quarries and mines of the Northwest Magnesite Co. near Chewelah, Wash., are large producers of magnesite for refractories. The deposits are extensive, some veins being hundreds of feet wide. The boundaries of the veins are indefinite, and the magnesite is interfingered with dolomite. Calcite, talc, serpentine, quartz and shale are disseminated through the deposits. The quality of the ore cannot be determined by appearance or by simple tests; chemical analysis is required. These features make it difficult to determine the boundaries or total quantity of workable ore.3 Only the higher-grade ores are mined, and a considerable proportion of the magnesite content is discarded by hand-sorting and log-washing methods. After calcining, the sorted product meets the specifications for usual refractory material, as it contains approximately 81 per cent MgO, 7 per cent SiO2, 6 per cent CaO and 6 per cent Fe203 + A1203. A sample of the reject material had the following analysis after calcining: 65.4 per cent MgO, 6.3 CaO, 21.1 SiO2 and 6.5 R2O3. The percentage reported as R2O3 includes compounds whose chlorides are acid-soluble but give precipitates in ammoniacal solution that are ignited and weighed. The more common constituents are Fe2O3, A1203 and TiO2. The sorted magnesite product is not the best grade. even for refractory purposes, and leaves much to be desired for a reduction process. The rejected material contains a large proportion of magnesite. It is apparent that an effective concentration method would yield a higher recovery and a better product. In general, the refractory grade is not suitably pure for other uses. Recent developments in the production of

Jan 1, 1943

By H. Wilton Clark

"IntroductionTHE CANADIAN use of short-period delay-action detonators in coal shooting has been carried on for about two years, mostly in British Columbia, and this is a preliminary report on progress to date. Entire coal faces are being shot successfully with one application of electric current.The electric blasting cap came into use about the beginning of the century. The low-tension electric cap, the type most commonly used, consists essentially of a metal tube containing an initiating explosive. Two wires lead into the end of the cap and are connected inside the cap with a short 'bridge wire .of very high resistance. This •bridge wire is buried in a match-head of a readily ignitable powder composition. When current goes through the lead wires, the bridge wire becomes extremely hot and sets• off the match-head which, in turn, ignites the powder just below it in the tube. This explosion sets off the main charge of explosive. Thus, by connecting a group of caps to a source of electric .current, a number of charges could be exploded simultaneously. This immediately became of great value to mine and quarry operators. Explosives could now be used to break up large masses of material at one time, as against the older, slow method .of blasting small amounts with single charges.The next step was to develop a delay detonator which would allow charges of explosive to fire one after the other, one or two seconds apart. Such detonators allowed first blasts to relieve some rock burden before the next charge fired, giving the second charge an easier task in forcing out more rock, and so on for the rest of the series of charges."

Jan 1, 1953

By Andrew Abraham

Continual improvement is the key to success in today?s competitive global arena. You require tools that promote dynamic quality management, deliver cost savings to your bottom line and provide the foundation for process improvement. ISO 9000 and ISO 14001 management systems and OHSAS 18001 are those tools. Since their inception, the benefits of ISO management systems and registration have been touted as the way to achieve better bottom line results. This spurred many companies to embark upon the road to registration with high expectations. The potential benefits of ISO management system implementation and third party registration that are commonly provided by Registrars and consultants are compelling adverts for success. ISO 9001:2000 is promoted as an essential tool for systematic efficient and affective management controls that provide a framework for analyzing errors and improving a company?s internal operational processes. It also allows companies to comply with internal and external customer and regulatory requirements thus continually improving the quality of their products and services and thus enhancing customer satisfaction. This, in turn, leads to greater customer recognition and increased market share. ISO 14001 implementation and registration may assist your company in reducing its litigation risk, raw material and energy consumption, waste disposal costs and possibly non tariff trade barriers. It should also provide continuity of operations, enhanced operational efficiency and consequently a competitive advantage in the market place. Registration of your ISO 14001 management system should also improve your company?s profile and credibility leading to regulatory flexibility and improved community goodwill. All of this will be a catalyst for improved morale and pride by employees and management alike. Likewise, OHSAS 18001, the latest certification specification for Occupational Health and Safety Management Systems supports Occupational Health and Safety legal and regulatory obligations and compliance, manages & controls Occupational Health and Safety risks and improves performance, safety culture and efficiency. OHSAS 18001 is an audit certification, and not a legislative requirement. Suggested benefits of implementation include reduction in accidents, lost time and associated liability and insurance costs and improving reputation and credibility through demonstrated commitment to health and safety. The foregoing increases access to new customers, joint venture partners and financing The skeptics say these are motherhood statements and lack corroborative evidence, i.e., ?Show me the money!?

May 1, 2003

By A Bird

Gravity gold circuits have an important part in reducing operating costs and maximising recovery in modern gold plants where coarse gold is present in the ore body. The design and selection of equipment has a vital part in the overall effectiveness of the gravity gold circuit. There are a number of associated factors that need to be taken into consideration when selecting the most suitable equipment for gravity gold circuits including occupational health and safety, security and labour requirements. This paper is a case study that outlines the validation behind the design, selection of equipment and subsequent benefits of the recent upgrade of the gravity circuit at the Marvel Loch gold mine with the main component to the upgrade being the installation of a Consep Acacia Reactor. Since the upgrade there has been a significant increase in gravity recovery, which has reduced the amount of gold reporting to the leaching circuit therefore reducing reagent consumption and overall operating costs. There has been an increase in overall gold recovery and therefore increased gold production. The benefits to occupational health and safety include the removal of toxic gases being released in the gold room during the previous practice of gold concentrate calcination and the removal of the arsenic and nickel from the gold doré, which poses a risk during refining. Manual handling and security has improved as there is no physical contact with gold concentrates during the automated process using the Acacia reactor. The total time for gold room personnel working in the gold room has also reduced from around six hours per day to less than one hour per day.With the continuing development of low grade gold deposits, there will be an increase in demand for low capital, high recovery gravity gold circuits that can significantly increase overall production while reducing operating costs.

Aug 8, 2011

By Thomas Deschamps

The main objective of this work was to investigate the effect of binder addition on the geochemical properties of surface paste tailings. Nine leaching columns were prepared following different disposal configurations: non-cemented, non-cemented with cemented layers, and cemented .The proportions of cement by total paste weight were between 0 and 1% for a whole column; the maximum amount of cement in a single layer was 2%. The material used was a sulphide-rich tailings coming from a Canadian hard rock polymetallic mine. The columns have been flushed once a week during 30 weeks. The recovered leachates were analyzed for pH, Eh, conductivity, and a wide range of elements. The drainage rates were monitored during the initial disposal as well as after each column flush. The results show that binder addition can improve the whole environmental performances for some of the studied disposal configurations. The disposal configuration consisting of intercalation of three 2% cement paste layers within non-cemented layers (0.66 wt.% of cement for the whole column), and the configuration with 2% cement added only in the two bottom layers (0,44 wt.% of cement for the whole column), showed the best results in terms of water quality and drainage rates.

May 1, 2007

By M. Samuels

"Block and panel cave mining methods are being increasingly proposed as they allow massive, deeply situated ore-bodies to be mined economically at a time when discovery rates of significant near surface deposits are declining. Cave-to-Mill, a mine-to-mill approach for block cave mines, defines ore block models with respect to both mine and mill performance. A cave-to-mill study for the New Afton block caving operation focused on identifying value opportunities related to fragmentation and sorting for a future lift as development progresses deeper. The lack of selectivity and potential for dilution associated with the block cave mining method results in many caving operations having to mine and either dispose of or process material that is below cut-off grade at certain stages of production. As an outcome of the study, the variation in cave fragmentation size and the sensitivity of mill throughput to feed size was examined. To evaluate the amenability of bulk and particle sorting, Prompt Gamma Neutron Activation Analysis and X-Ray Fluorescence sensors were tested. A conceptual bulk and particle sorting flowsheet is presented. Grade data from regular underground sampling was used to predict the performance of a bulk sorter. A heterogeneity analysis of material reporting to drawpoints helped to identify areas of the extraction level that experience high variations in grade and thereby stand to benefit from dedicated ore and waste material handling systems should underground bulk sorting be implemented.INTRODUCTION Block and panel caving are low-cost underground mass mining methods that are suitable for excavating weak, massive and steeply-dipping orebodies. Accordingly, cave mining methods have predominantly been used to mine copper porphyry and diamond bearing deposits. Due to the increasing depletion of near-surface orebodies, it is expected that block and panel caving methods will be increasingly used in the copper-gold industry. Their implementation holds significant potential to extend the life of existing open-pit operations by providing an economic means to access ore located at depths that are beyond the practical limits of open-pit mining. Examples of cave mines proposed for British Columbia include the Kwanika Mine, Red Chris underground, Kemess Underground project, and the Iron Cap and Mitchell cave mines of the Kerr-Sulphurets and Mitchell (KSM) project (Moose Mountain Technical Services, 2017; Imperial Metals, 2017; Golder Associates, 2017; Tetra Tech, 2016)."

Jan 1, 2018

By C. Babcock

When a flat-lying mineral deposit such as a coal seam is mined, some coal is left to separate the roof and floor so that miners and machines can safely mine the seam. The mineral left is in the form of mine pillars. If too much is taken, support problems and unsafe working conditions result. If too little is taken, the mining operation will be unable to compete economically with other mines. Therefore, one must select pillars of the optimal size to maximize mining recovery and minimize risk to miners and equipment. Later, if the pillar support is no longer needed and the surface can be allowed to subside the thickness of the coal seam, then the pillar support can be removed to increase mining recovery, and initial pillar size is not as critical. Only a small part of the information available on pillar design can be presented here. The purpose of this report is to summarize some of the work done by the U.S. Bureau of Mines and others on pillar design, which is reported in the literature from 1833 to the present, and present rock mechanics test results that aid in understanding mine pillar behavior. The variable names used are those of the authors cited. This results in multiple use of some variable names.

Jan 1, 1994

By L. W. Bahney

(Cleveland Meeting, October, 1912.) THE photograph, Fig. 1, shows the apparatus a little less than half size, consisting of a filtering-flask fitted with rubber stopper, through which passes a bent glass tube, and an extraction-thimble fitted with rubber stopper through which passes a glass tube of 0.2-5-inch bore. Both tubes are connected by a short piece of rubber tubing. A section of a thimble is shown in the photograph; the tube extends to within in. of the tapered end. The object of using the thimble is to remove the acid from the beaker after all the copper has been precipitated. Time is saved, the copper is not exposed to the acid alone, and there are none of the losses attending ordinary filtration. I have accomplished these results by means of a piece of perforated platinum fastened in the end of a 0.25-in. bore glass tube and a filter-mat of asbestos, but after my supply of proper length fiber became exhausted I could not replenish it even after purchasing 14 lots from four different dealers. The above apparatus may be used to remove at least seven-eighths of a supernatant liquid from a settled precipitate without disturbing the latter. The application of the thimble is best shown by partly out-lining the assay for copper, as follows: Dilute the acid solution of copper and other sulphates to 150 cc. in a 200-cc. Jena beaker, place on a hot plate, add 2 drops of concentrated HC1, then place a strip of aluminum in the beaker (this may be bent or straight, as desired). Connect the apparatus, as shown in Fig. 1, with a filter-pump having a, strong suction. When the copper is precipitated, remove the beaker from tine hot plate and insert the extraction-thimble alongside the strip of aluminum. The acid solution will be drawn through the porous tube. Wish the upper end of the aluminum strip with a jet of hot water, wash down the sides of

Nov 1, 1912