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By Mariam Melashvili, Chris Fleming, Rene Jackman

"A study was undertaken to develop a flowsheet for treating a high grade silver/gold ore body by cyanidation, CCD and Merrill Crowe precipitation, with tailings detoxification to meet stringent limits for cyanide and thiocyanate concentrations in the final discharge. In one of the flowsheets examined, free cyanide in the final leach liquor was recycled to the mill where it was hoped that it might react beneficially with fresh ore to reduce the cyanide makeup by an equivalent amount. However, laboratory results suggested that most of the cyanide recycled to the mill reacted with iron forming a soluble ferrocyanide complex. This would have had a negative impact on cost, not only because the cyanide recycled to the mill could not be claimed as a credit, but also because the presence of ferrocyanide in the final tailings meant copper sulphate had to be added to the SO2/air cyanide detoxification process to precipitate the iron cyanide. In an alternative flowsheet that was examined, recycling of cyanide to the mill was abandoned, and this allowed the milled slurry to be pre-aerated in a single mixing tank before cyanidation. During alkaline pre-aeration, any ferrous ions that are generated in the mill are oxidized to ferric ions, which, unlike ferrous ions, are inert in cyanide solution. This significantly reduced cyanide consumption and lowered costs in the detoxification of the tailings. In this paper, the cost implication of these two circuits is compared.INTRODUCTIONMost of the gold that is produced by hydrometallurgical processes is recovered by cyanide leaching. The reason for that lies in the superior thermodynamic stability of the gold cyanide complex and the favorable economics this confers on the process (Fleming, 2014). However, the gold in most of the gold-bearing ores processed is a very minor component, and therefore, the leaching response is often dominated by that of the host minerals.Many gold ores treated by the cyanidation process contain iron minerals. Cyanide reacts readily with some iron minerals forming ferrocyanide as one of the compounds. Ferrocyanide is stable when solutions are exposed to air and does not retard the dissolution of gold. In fact, it has been claimed that potassium ferrocyanide in the presence of oxygen is a weak solvent for gold (Hedley & Tabachnick, 1958). The main problem with the formation of this compound is the consumption of cyanide. Ferrocyanide contains six molecules of cyanide for each atom of iron {Fe(CN)64-} so the wasteful consumption of cyanide is greatly increased by its presence. Thiocyanate is also often produced during the reaction of certain iron sulphide minerals with cyanide, and contributes to the wasteful consumption of cyanide. In addition, environmental regulations in certain countries demand that thiocyanate be removed from the tailings, and this adds significantly to the cost of tailings treatment."

Jan 1, 2016

By Pietro Navarra

Cooling of critical furnace equipment to produce freeze lining is of particular interest as many metallurgical processes are intensified while attempting to prolong campaign life. This idea was combined with high-capacity heat pipes in the design of a copper slag launder. The design methodology of the launder and heat pipe cooling system is examined in this paper. With reliable material properties and operating conditions, CFD was used to quantitatively predict the skull thickness and the corresponding heat load applied to the slag launder. The operational limitations of heat pipe technology were considered and an appropriate cooling system was incorporated into the model. Parametric modeling was used to simulate several launder and heat pipe configurations. The generated results were then used to optimize the design of the launder, which was built and tested in August of 2005. The experimental performance of the cooling system is evaluated and compared with the model results.

Jan 1, 2006

By O. S. True, C. H. Herty

It is well known that manganese will desulfurize molten iron through the formation of manganese sulfide, which, being only slightly soluble in the metal, rises to and enters the slag where it remains as such (mixer slag) or is converted partly or wholly to calcium sulfide (basic open-hearth slag). The reaction by which manganese sulfide is formed is Mn + FeS = MnS + Fe. This reaction is generally considered to be reversible, although McCancel states that such is not the case. The experimental work was carried out to confirm the results of Roh12 as to the reversibility of the reaction. Six trials are herein considered. In each 200 gm. of electrolytic iron free from manganese and sulfur were melted in a pure zirconia crucible in a carbon resistance electric furnace. To the melted iron in runs 2, 3 and 4 metallic manganese and manganese sulfide were added to give 1, 2 and 4 per cent. manganese respectively, with 0.50 per cent, of sulfur in each. In runs 5, 6 and 7, metallic manganese and iron sulfide to give the same proportions of. manganese and sulphur as in runs 2, 3 and 4, were added. The melts were held molten in the furnace for an hour and then cast into small ingots. The results of these melts are as follows: Ingredients of Charge Analysis or InRot ManOaneBe, SulfUR, ManOaneBe. SULFUR, Run Material Charged PeR Cent. PeR Cent. Per Cent. PeR Cent. 2 MnS + Mn............... 1.0 0.452 0.636 0.367 3 MnS + Mn............... 2.0 0.50 1.57 0.310 4 MnS + Mn............... 4.0 0.50 2.81 0.230 5 FeS + Mn................ 1.0 0.50 0.647 0.397 6 FeS + Mn................ 2.0 0.50 1.45 0.374 7 FeS + Mn................ 4.0 0.50 2.73 0.195

Jan 1, 1925

By J. M. Pachter

Property sales and acquisition are prominent features of today's coal industry. Before selling or buying a site, a company must know the property's economic net worth. Any evaluation of economic net worth would be inadequate if determinations of a property's environmental liabilities and costs were omitted. Short-term and long-term costs can result from liabilities related to reclamation, water treatment, subsidence, and hazardous waste disposal. A structured, interdisciplinary approach can aid in the delineation, summation, and scheduling of these costs and liabilities.

Jan 1, 1986

By Marco Antonio Falquete

A novel spark-generating tube, containing no explosive in its composition, is evaluated and compared with conventional shock tube. The product overcomes a series of problems found in conventional shock tube technology. The technologic development had multiple aims: to avoid transport and storage costs associated with explosives, to be more environmental-friendly, to obtain a high-velocity, hightemperature spark whose thermal performance could overcome poor heat transfer to the delay charge by gas convection at shock tube end, and that could cross barriers such as tight kinks, close knots, cuts, extreme environmental conditions, and long-term hot explosive emulsion exposure. The product of this development is evaluated and compared with conventional shock tube in a series of aspects and performance tests, including spark thermal performance (ignition of tougher, low sensitive delay columns, spark gap test), propagation in critical conditions of use (kinks, knots, cuttings), extreme environmental conditions (long-term exposure to hot emulsion explosive, extremes of cold, heat, moisture, underwater exposure), safety features (ignition through flame, open fire), and security against terrorist acts. Due to the safe characteristics of the tube, a series of demonstrations will be performed during the presentation. In nearly all tests and aspects, the spark-generating tube exceeds conventional shock tube, and in the rest equals shock tube, making it a more reliable alternative.

Jan 1, 2005

By J. A. Batista-Rodríguez, F. J. López-Saucedo

A new methodology is proposed for the preliminary assessment of coal blend quality, which combines gamma-ray spectrometry and magnetic susceptibility measurements. This methodology does not replacethe conventional methods of assessing coal quality, but can help to control the quality of the coal blending process, as well as reduce the amount of proximate analysis required and the determinations of calorific value and total sulphur. The methodology includes five steps. First, both geophysical measurements aremade with handheld equipment. The samples with magnetic susceptibility ≤ 1 × 10-3 SI are then separated. From these samples, those with total gamma intensity ≤ 20 nGy/h are separated. These samples have a higher calorific value and lower ash and moisture content. To separate the samples of better quality and with lower total sulphur content, in step four the samples with total gamma intensity > 20 nGy/h are separated from those with magnetic susceptibility ≤ 1 × 10-3 SI. This last group includes samplesthat will cause less total sulphur-related environmental pollution. In the final step, the quality of both radioactive sample groups is corroborated from the calorific value measurement and ash, moisture, andtotal sulphur contents in at least 10% of all coal blend samples from both groups. In this study, more than1000 samples were analysed with this new methodology, and the results are remarkably similar to the ASTM classification of the Sabinas Basin coals.

Apr 2, 2022

Use of Impact Crushing for the Enrichment of Low-Grade Colemanite Concentrates This paper presents the results of an investigation into the enrichment of -25+3 mm low grade colemanite concentrates exploiting differential breakage properties of various minerals associated within the ore sample. Boric acid production process requires concentrates containing at least 42 per cent B2O3, hence concentrates below this grade either stock piled or further treated to increase B2O3 content. The low grade concentrate is obtained as a by-product of Emet Colemanite Concentrator and the B2O3 content may vary around 37 - 38 per cent. The sample used in the experimental studies assays 37.52 per cent B2O3. In the experimental procedures the sample was crushed using a laboratory size horizontal shaft impact crusher and the products were classified into different size fractions by screening so that any improvement by size could be observed. In the tests the effects of parameters such as the revolution speed of the crusher and heat treatment was checked. Test results prove that boron contents show a tendency to increase as the particle size decreases. As a result of the experimental studies a colemanite concentrate assaying 42.13 per cent B2O3 is produced with 76.55  per cent recovery. Arsenic content is seen to change inversely proportional with the B2O3 content of the concentrate such that it drops from 1016 ppm from the as received sample to 676 ppm in the concentrate.

Sep 13, 2010

By S Jeon, K Haga, A Shibayama, L Godirilwe, C B. Tabelin, I Park, A Buronov

Open pit/underground mining and mineral processing can generate large amounts of waste that contains sulfide-type minerals such as pyrite (FeS2), and it can produce acid mine drainage (AMD) when exposed to oxygen and water. This highly acidic wastewater contains heavy metals and this is one of the most serious environmental issues after mining operations taking several decades to completely remediate with high costs. For green mining, the present study focused on the development of a recovery technique for gold from an eco-friendly solvent, thiosulfate, by utilising mine wastes, tailings. First, gold recovery experiments in ammonium thiosulfate system were carried out using model samples representatively present in tailings (ie SiO2, Al2O3, FeS2, Fe2O3, and Fe3O4). The results showed that gold recoveries were negligible when using Al2O3, SiO2, Fe2O3, and Fe3O4 while pyrite shows a recovery efficiency three-fold as high as the maximum recovery efficiency when using others. Subsequently, the study also addressed the identification of the recovery mechanisms and post-recover process. The results indicate that tailings can be sufficiently utilised for gold recovery in the hydrometallurgy process and a further sustainable process can be established.

Sep 1, 2024

By C. M. Trigg

"PEC uranium deposit, 100 km southwest of Coppermine, was discovered by diamond drilling in 1976 following ground and airborne geophysical surveys, geological mapping, geochemical surveys and boulder prospecting that were initiated as the result of the discovery of uraniferous sandstone outcrops in 1969 by geologists of Aquitaine Company of Canada Ltd.The deposit is within gently northeasterly dipping medium to coarse-grained Unit 11 quartzitic sandstone of Neohelikian Dismal Lakes Group. The deposit is from 10 m to 70 m above the base of the sandstone which, beneath the deposit, non-conformably overlies weathered and altered Aphebian granite of Great Bear Batholith or unconformably overlies rubble breccia, conglomerate and medium-grained Unit 8 sandstone of Paleohelikian Hornby Bay Group.A northeasterly trending, 500 m wide graben, bounded by the steeply dipping northeasterly striking Imperial and Aquitaine faults, contains the deposit which is 1.1 km long by up to 340 m wide. The deposit overlies a post-Hornby Bay Group, pre-Dismal Lakes Group topographic high. Uranium principally occurs as stratabound, often peneconcordant bands; a lesser amount of uranium is fracture controlled.Stratabound uranium exists in gently easterly dipping, subparallel bands ranging in thickness from 1.5 m to 6.0 m. Pitchblende, coffinite and sooty black uranium oxides occur as interstitial disseminations and grain coatings subparallel to bedding; they also replace pyrite and form narrow veinlets.Small amounts of torbernite, carnotite and cuprosklodowskite coat pitchblende and are disseminated within mineralized bands. Pyrite, marcasite, chalcopyrite, chalcocite, cove/lite, digenite, azurite, malachite, arsenopyrite, glaucodot, niccolite, bravoite (? ), goethite, hematite and magnetite as well as quartz, barite, calcite, biotite, chlorite and clay minerals exist within the mineralized bands. Uranium grades of 0.1 % to 0.3 % U3O8 over thicknesses of 1 m to 2 m are typical.A steeply dipping fracture zone near the Imperial fault contains sooty black uranium oxides, torbernite, carnotite and cuprosklowdowskite as well as sulphides, cobalt arsenides, quartz and carbonate. The zone may reach 70 m long; its true width is about 1.9 m. Uranium grades range up to 5.18% up, over0.9m.Pitchblende and coffinite from PEC deposit have been dated at 1.08 ± .10 Ga with possible remobilization at .45 Ga. The deposit contains an inferred 3,150 tonnes of U3O8."

Jan 1, 1986

By Rudolf Kudlich

The problem of eliminating the hazards of hoisting in mines has been with us since the industry passed its earliest stages, when coal and ore could be won from surface working and tunnels. At first, safety devices were introduced as means of decreasing the amount of labor necessary to produce the mineral, or to increase the amount that could be obtained from a certain opening. For example, crude ladders replaced winding footpaths or steps cut into the walls of the pit or shaft, decreasing the distance it was necessary to carry the product to the surface and eliminating, to some extent, the hazard introduced by the precarious footing afforded by such paths and steps. Latcr, instead of being carried from the pit on the backs of men, the product was hoisted in buckets or baskets on a rope passing over a pulley or around a windlass. Thus the capacity of the pit was considerably increased, and the hazard due to making many trips up and down the shaft was eliminated. From such beginnings accident prevention in mine hoisting advanced with the importance of the industry, until at present hoisting compares favorably with other branches of the mining industry with reference to safe practices. As in the other branches, this progress has been accomplished along two general lines—by legislation, rules, and education, and by the installation of safety devices. Though due credit must be given the former, it is not so satisfactory as the latter. Legislation, to be effective, requires a policing and inspection system to guarantee enforcement, such as we cannot hope to inaugurate. Rules may be satisfactory so long as the employer controls the labor market, and dismissal or some other form of discipline is an effective means of insuring obedience to the rules; but where labor demand is in excess of the supply, or where labor organizations are in control, such measures are not effective. An educational campaign will draw attention to hazards and may awaken a desire to avoid accidents by exercising the proper care. It is necessary, therefore, to place our reliance mainly on mechanical devices to prevent accidents, and to depend as little as possible on laws,

Jan 1, 1923

By Rudolf Kudlich

The problem of eliminating the hazards of hoisting in mines has been with us since the industry passed its earliest stages, when coal and ore could be won from surface working and tunnels. At first, safety devices were introduced as means of decreasing the amount of labor necessary to produce the mineral, or to increase the amount that could be obtained from a certain opening. For example, crude ladders replaced winding footpaths or steps cut into the walls of the pit or shaft, decreasing the distance it was necessary to carry the product to the surface and eliminating, to some extent, the hazard introduced by the precarious footing afforded by such paths and steps. Latcr, instead of being carried from the pit on the backs of men, the product was hoisted in buckets or baskets on a rope passing over a pulley or around a windlass. Thus the capacity of the pit was considerably increased, and the hazard due to making many trips up and down the shaft was eliminated. From such beginnings accident prevention in mine hoisting advanced with the importance of the industry, until at present hoisting compares favorably with other branches of the mining industry with reference to safe practices. As in the other branches, this progress has been accomplished along two general lines—by legislation, rules, and education, and by the installation of safety devices. Though due credit must be given the former, it is not so satisfactory as the latter. Legislation, to be effective, requires a policing and inspection system to guarantee enforcement, such as we cannot hope to inaugurate. Rules may be satisfactory so long as the employer controls the labor market, and dismissal or some other form of discipline is an effective means of insuring obedience to the rules; but where labor demand is in excess of the supply, or where labor organizations are in control, such measures are not effective. An educational campaign will draw attention to hazards and may awaken a desire to avoid accidents by exercising the proper care. It is necessary, therefore, to place our reliance mainly on mechanical devices to prevent accidents, and to depend as little as possible on laws,

Jan 1, 1923

By Leonard Harris, Rosa H. Harris

"It is becoming increasingly obvious to mining and exploration companies, when operating overseas, that their activities should not conflict with the interests of the surrounding communities. The world's largest gold heap leach operation, Minera Yanacocha has installed community assistance programs in Peru concerned with health, education, electrification, agricultural assistance and road improvements. The programs are conducted with joint venture partners consisting of the company, the community people, the contractors and suppliers, government and non government organizations and a ladies association composed of the employees' wives and female staff.IntroductionThe l 990's is the decade highlighted by enormous investments in mineral exploration and mining projects in the developing world. Figure 1 illustrates the distribution of exploration spending by mining companies in 1995."

Jan 1, 1997

By S. Paul Singh

Rock is a heterogeneous medium, a fact often ignored during blast design. In reality, the rock mass characteristics have a controlling influence on the outcome of a blast. If the optimum blasting results are to be achieved over a range of rock mass conditions, it is prudent to understand the role played by the geological features of a rock mass in producing overbreak around the periphery of an excavation. The field work was performed in a drift located in a geologically dominant rock mass and also included observations and assessment of blast damage at several operating mines. The blast damage was estimated on the basis of the half cast factor. The effect of rock strength, rock quality designation, orientation , frequency and aperture of joints, number and thickness of layers and hydrogeological conditions on blast damage has been described.

Jan 1, 2000

By Augustus Locke

The paper has two parts: the field argument and the microscopic argument. The field argument indicates that either the deep chalcocite has replacea sulfide of extraordinarily easy replaceability or it is hypogene. The microscopic argumentindicates that this sulfide, if it existed, must have been hypogene bornite; that this bornite must so excel known bornite in replaceability as to constitute a new variety; that the existence of this new variety is wholly unsupported by evidence; and that the deep Butte chalcocite is therefore hypogene. IN 1900, when the public first heard of "secondary enrichment" the Butte chalcocite seemed clearly supergene. Mining, through successive regions of leached capping, bonanza sulfide, and sulfide less rich, had paused at lean bottom levels which, though chalcocite-bearing, suggested the waning of copper. The deepest chalcocite, at 1500 ft., was yet too shallow to disturb the supergene hypothesis. But, with deeper mining, lean levels yielded to fat ones. Here and there they showed, within the same vein, an inverted distribution, chalcopyrite above and chalcocite below. To the miners' astonishment, they carried chalcocite, unaffected by cross faults, on an independent course to the depth of 2500 ft. The pattern of ore distribution, hitherto revealed only in fragments, was now appearing in perspective, and, in the depth of the chalcocite, was beginning to make Butte differ from all districts where the chalcocite is indubitably supergene. The idea of hypogene origin at length came into favor; by 1913, when Sales1 advocated this origin, the argument had become a powerful one. This argument rested on field evidence; but, already, there was beginning to emerge microscopic evidence not included in this arugment. Especially pertinent was the proof that (1) The deep chalcocite lies in-mixed with bornite, often microscopic; (2) the deep chalcocite often replaces bornite; and (3) replacement of bornite by chalcocite is accom-

Jan 3, 1924

Zircon producers are often required to lower impurity levels to meet customer specifications as most of these impurities adversely effect the end use. For example, zircon sand in foundry applications needs to be free of any surface impurities (eg clay, gibbsite and goethite) which otherwise could cause sintering or fusion between cast metal and the zircon mould. Light coloured ceramic grade zircon, typically low in iron oxides and titania, is preferred in ceramics applications. The presence of impurities in zircon sands is a problem in the magnetic and electrostatic separation stages of their concentration. Previous work (Aral et al, 1997) showed that impurities in zircon sand concentrates occurred in the form of stains on grain surfaces, discrete gangue minerals and in the mineral grains. The gangue minerals such as rutile, garnet, xenotime, monazite, spinels, ilmenite, and aluminium silicates in the concentrate in some cases (Aral et al, 1997) contribute significantly to the impurity content of the concentrate. A careful hand picking of the impurity grains from a Western Australian zircon sand showed that the impurity content of the concentrate could be lowered significantly as a result of the removal of the surface stained grains and gangue minerals. The hand-picked ædirtyÆ fraction, comprised about eight per cent by weight of the concentrate, and consisted mostly of iron-oxide/hydroxide-stained zircon, silica grains and rutile, and occasionally ilmenite, monazite, gahnite and other (unidentified) coloured grains. The chemical analysis of the remaining æcleanÆ fraction showed that the major impurities (Fe and Ti) were reduced significantly (Table 1). The reduction in the uranium content was about 30 per cent. The other measured impurities, P, Mg, Al, Th, and Ca remained mostly unchanged as shown in Table 1. The contribution of ilmenite to the overall iron content of æcleanÆ zircon was insignificant as it comprised about 0.1 per cent weight of the æcleanÆ zircon. Rutile was the major source of titanium contamination (about one per cent by weight). Recent work (Aral and Calle, 1999) on zircon sand concentrates from the Murray Basin showed that, under the electron microprobe, the bulk of the impurities occurred in the dark back scattered electron (BSE) regions of the zoned zircon grains. The impurity contents of the light BSE regions and unblemished (unzoned) grains were low. The impurity content in zoned zircons has been investigated on single, relatively large, crystals (McLaren, Gerald and Williams, 1994; Smith et al, 1991; Romans, Brown and White, 1975). These studies generally aimed to investigate the incorporation of U in zircon as zircon is used extensively for U-Pb geochronology owing to its relatively high content of uranium. Smith et al (1991) identified a striking difference in composition between the areas of lower (dark BSE) and higher (light BSE) average atomic number in the zoned zircon. Their work showed in particular that the zones of lower atomic numbers contained very significant amounts of trace metals such as Y, REE, Nb, U, Al, Fe, Ca, Ba and, by difference H2O. In this paper the impurity distribution in zoned zircon grains was investigated on statistically significant number of grains before and after a leach treatment. The distribution of impurities was also measured in unblemished areas to compare the results.

Jan 1, 2000

Study on the Reasonable Mineral Processing Technology of a Cu-Mo Ore

Sep 13, 2010

By Menne D. M, Ladyman R

Camel Creek Mine operated by Lynch Mining Pty Ltd is located 65 km west of inghain and 160 kin north west of Townsville in North Queensland. (Figure 1). The Camel Creek gold mineralisation was discovered in mid-1987 in an area previously known for its tin workings. Mining commenced in January 1989 after an oxide reserve of 1.2 million tonnes @ 2.1 g/t gold had been delineated. The first gold pour was in May 1989. The mine developed into a medium sized open cut operation, with full site facilities. The initial mining rate was 500 000 tpa utilising run-of-mine (ROM) heap leaching to recover the gold. By the end of 1990 the wider orehodies were mined and efforts were focused on the narrow orebodies, of width typically less than three metres. Smaller earth moving equipment was employed to selectively mine the narrow orehodies leading to a decrease in the mining rate to 200 000 tpa. The Red Gold deposit (80 000 tonnes « 3.0 g/t to 18 in vertical) was mined in mid-1990 and ore was hauled approximately 8 kin to the Camel Creek heap leach operation. The ore was hauled in a dedicated campaign after being stockpiled. The cost of dedicated ore haulage ensured that only the more exceptional of the remote orebodies could be mined in this manner. The Golden Cup orebody, some 20 km from Camel Creek, was too distant to permit ore haulage. However, this prospect contained sufficient gold (initially 90 000 tonnes @ 3.0 g/t) to warrant the installation of a small mobile adsorption facility. Construction of the second heap leach operation commenced in late-1990. Gold production commenced in early April 1991. Mining of several smaller orebodies remote to the second heap leach facility has been undertaken and further mining is planned. This paper outlines heap leaching practice at Camel Creek Mine and summarises the success of the Golden Cup heap leach.

Jan 1, 1994

By Khaled M. Mohamed, Michael M. Murphy, Ted M. Klemetti, Heather E. Lawson

"The design of rib support systems in U.S. coal mines is based primarily on local practices and experience. A better understanding of current rib support practices in U.S. coal mines is crucial for developing a sound engineering rib support design tool.The objective of this paper is to analyze the current practices of rib control in US coal mines. Twenty (20) underground coal mines were studied representing various coal basins, coal seams, geology, loading conditions, and rib control strategies.The key findings are: (1) any rib design guideline or tool should take into account external rib support as well as internal bolting; (2) rib bolts on their own cannot contain rib spall, especially in soft ribs subjected to significant load—external rib control devices such as mesh are required in such cases to contain rib sloughing; (3) the majority of the studied mines follow the overburden depth and entry height thresholds recommended by the Program Information Bulletin (PIB) 11-29; (4) potential rib instability occurred when certain geological features prevailed—these include draw slate and/or bone coal near the rib/roof line, claystone partings, and soft coal bench overlain by rock strata; (5) 47% of the studied rib spall was classified as blocky—this could indicate a high potential of rib hazards; and (6) rib injury rates of the studied mines for the last three years emphasize the need for more rib control management for mines operating at overburden depths between 500 ft and 1,000 ft.IntroductionEfforts to improve the stability of underground mine ribs have continued for decades. These efforts by mining professionals around the world persisted throughout the 1990s and 2000s as well, though much of this work focused on rib support methodology and selection of proper support for a specific location. Despite these efforts, it is unfortunate that there has been a continual occurrence of rib-related fatalities in underground coal mines. While overall improvements in rib safety have been achieved, the average fatality rate is still about 1.3 fatalities per year over the last 18 years (Jones, 2014).To manage and control the rib hazard, rib support is used in U.S. coal mines. The current rib support methods for U.S. mines fall into two main categories: rib control based on intrinsic support systems and those based on external support systems. Intrinsic or rib-bolt support requires a roof bolting machine that can install a bolt fixture into the rib. Furthermore, the appropriate level of support needed to control the ribs with an intrinsic-based support system as well as the design parameters for those support systems for various mine conditions have not been established in U.S. coal mines. Furthermore, surface control systems are often an integral part of these support systems and must be considered in the system design."

Jan 1, 2015

By Amilton Feliciano Castro

In this work, activated carbons (ACs) were prepared using Sewage sludge as starting material. The ACs were pyrolyzed in an inert atmosphere at 550 °C after independent chemical treatment with ZnCl2 or KOH with ratio of 1:1. The materials were used as adsorbents for removal of Cu(II) from aqueous solutions. The adsorbents were characterized using adsorption/desorption nitrogen isotherm, SEM (scanning electron microscopy). The results showed that copper removal could be linked to surface pH and compounds present on the surface of adsorbent. KOH-treated adsorbent has smaller surface area (186 m2 g-1) than ZnCl2-treated adsorbent (192 m2 g-1), however, KOH-treated adsorbent exhibited higher removal of Cu(II). The performance of KOH-treated adsorbent could be attributed to the presence of basic groups on the adsorbent surface. Liu isotherm model gave the best description of the equilibrium data. From the Liu isotherm, the maximum adsorption capacities at 298 K for KOH-treated, ZnCl2-treated and sewage sludge (without inorganic activation) adsorbents are 31.85, 19.79 and 3.513 mg g-1, respectively. The best desorption experiment was obtained using 1.50 mol L-1 of HNO3 as eluent. Recoveries of Cu(II) from Cu(II)-loaded adsorbents are 98.9% (ZnCl2-treated) and 95.5% (KOH-treated.

Aug 17, 2017

By W. M. Robertson

The formation of grain boundary grooves on surfaces of poly crystalline samples of cobalt and iron immersed in liquid lead has been studied. The grooves form by volume diffusion of the solutes cobalt and iron in the liquid. The diffusion coefficients of the solutes in liquid lead are derived from the measured rate of grooving. The diffusion coefficients are described by the relation D = Do exp (-Q/RT), with, for cobalt, Do = 4.6 x 10-4 sq cm per sec and Q = 5300 ± 800 cal per mole, and for iron, DO = 4.9 x 10-3 sq cm per sec and Q = 10,500 ± 1500 cal per mole. LIQUID metal-solid metal interactions occur at solid-liquid interfaces. Interfacial energy provides a driving force to change the morphology of the interface. Mullins1,2 has derived expressions for the kinetics of interface morphology changes driven by capillarity. These expressions can be applied to an isothermal system of a solid in equilibrium with a liquid saturated with the solid. Surface profile changes can occur by volume diffusion of the solute in the liquid, by volume self-diffusion in the solid, and by interfacial diffusion at the liquid-solid interface. A groove will form at the intersection of a grain boundary with a solid-liquid interface, reducing the total interfacial free energy of the system. The solid-liquid interfacial energy ? must be greater than half the grain boundary energy of the solid ?6 for Mullins' calculations to apply. If ? is less than ?b/2, then the liquid penetrates the boundaries, separating the grains rather than forming grooves. Boundary penetration did not occur in the work described here. where CO is the equilibrium volume concentration of the solid in the liquid, Dv the volume diffusion coefficient of the solid in the liquid, ? the interfacial free energy of the solid-liquid interface, O the atomic volume of the solid crystal, k Boltzmann's constant and T the absolute temperature. Eqs. [1] and [2 ] also apply to grooving by volume self-diffusion in the solid,1 with CoODv = D Self, where DSelf is the volume self-diffusion coefficient of the solid. For a grooving mechanism of interfacial diffusion at the solid-liquid interface, the groove width is given by2 where CS is the interfacial concentration of the diffusing species, and DS is the interfacial diffusion coefficient. Eqs. [1] and [3] can be used to determine the mechanism of groove growth. A t1/3 dependence of the growth indicates volume diffusion and t1/4 indicates interfacial diffusion. In some cases, volume diffusion and interfacial diffusion both can contribute substantially to the grooving process, causing the time dependence to be intermediate between t 1/3 and t1/4.3 For these cases, the relative contributions of the two processes can be separated.4 However, in many cases, one process will be dominant, and the data can be analyzed on the basis of Eq. [1] or Eq. [3] alone. The time dependences for volume diffusion in a liquid and volume self-diffusion in a solid are the same. However, the self-diffusion contribution of the solid is usually negligible compared to volume diffusion in the liquid. After the grooving mechanism has been determined, Eq. [1] or Eq. [3 ] yields the kinetic parameter A or B. The kinetic parameter can be used to calculate values for the unknown quantities in the product CD?. Usually C is known or can be estimated. If ? is known, then D can be calculated. In a measurement of grain boundary grooving of copper in liquid lead,' the time dependence indicated volume diffusion in the liquid. The quantities Co, Dv, and ? were obtained from the literature, giving excellent agreement between the observed values of A and the values calculated from Eq. [2 ].5 In a study of the grooving of several refractory metals in liquid tin and liquid silver, A1len6 educed that grooves formed by volume diffusion in the liquid. In a study of nickel in a nickel sulfide melt, Steidel, Li, and spencer7 found volume diffusion grooving kinetics. Both Dv and ? were unknown, so they could not obtain either one separately, though they did obtain a reasonable value for the temperature dependence of the product Dv ?. Several methods have been used to obtain surface profiles. It can be done by sectioning through the interface7 or by chemically removing the liquid from the solid surface after solidification of the liquid.6 However, if the liquid dewets the solid on removing the solid from the melt, then the interface can be observed directly. This method was used previously' and was utilized also in the present study. EXPERIMENTAL PROCEDURE Lead of 99.999 pct purity was obtained from American Smelting and Refining Co. Cobalt sheet was obtained from Sherritt-Gordon Mines, Ltd., with a nominal purity of 99.9 pct, the principal impurities being nickel, iron, copper, carbon, and sulfur. The sheet was

Jan 1, 1969