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By P. T. Bissert, L. Yan, M. D. Klein

"Federal regulations require the installation of refuge alternatives (RAs) in underground coal mines. Mobile RAs have a limited ability to dissipate heat, and heat buildup can lead to a life-threatening condition as the RA internal air temperature and relative humidity increase. The National Institute for Occupational Safety and Health Office of Mine Safety and Health Research performed heat testing on a 10-person tent-type training RA and contracted ThermoAnalytics, Inc. to develop a validated thermal simulation model (TSM) of the tested RA. The TSM was used to examine the effects of the constant mine strata temperature assumption, initial mine air temperature (MAT), initial mine strata surface temperature (MSST), initial mine strata temperature at depth (MSTD), and mine strata thermal behavior on RA internal air temperature using 117 W of sensible heat input per simulated miner. For the studied RA, when the mine strata temperature was treated as a constant, the final predicted RA internal air temperature was 12.8°F lower than it was when the mine strata thermal behavior was included in the model. A 10°F increase in the initial MSST resulted in a 7.1°F increase in the final RA internal air temperature, whereas a 10°F increase in the initial MSTD yielded a 2.5°F increase in the final RA internal air temperature. INTRODUCTION Since 2008, the Mine Safety and Health Administration (MSHA) has required the installation of refuge alternatives (RAs) in underground coal mines (1). MSHA requires that RAs provide an environment with breathable air for entrapped miners for a 96-hour period. Heat buildup inside an occupied RA is a serious concern. Without a means to dissipate the heat and humidity generated by the occupants and the carbon dioxide scrubbing system, the temperature and humidity inside RAs could lead to severe discomfort or heat stress depending on the mine ambient temperature before and during occupation of the RA. NIOSH recommended an apparent temperature limit of 95°F in its 2007 report to Congress (2). In 30 CFR 7.504, MSHA has specified a maximum apparent temperature of 95°F inside an occupied RA (3). The apparent temperature is calculated using both air temperature and relative humidity (4). Calculations could be used to determine if the air within an occupied RA will reach the apparent temperature limit. In these calculations, in order to simplify the problem, the mine temperature could be assumed to remain constant. With this assumption, the effects of the mine strata thermal mass and thermal conductivity on the heat buildup in an RA are ignored. OMSHR research has shown that the mine air and strata temperatures surrounding an RA increase when subjected to the heat of an occupied RA (5). Thus, calculations performed with the assumption that the mine temperature is constant would under-predict the resulting dry-bulb temperature and apparent temperature for a given RA occupancy. The mine air temperature has been used as the basis for RA temperature calculations. However, research has not been used to confirm that the mine air temperature is the appropriate basis."

Jan 1, 2016

By Jae-Chun Lee, Sadia Ilyas

"The present study was carried out to develop an economically feasible and environmentally friendly technique to extract metals such as Al, Fe, Cu, Zn and Ni from electronic scrap by employing Penicillium chrysogenum strain KBS3. Studies were carried out in various leaching modes, including one-step leaching with combined growth and leaching phases (Mode 1), two-step leaching with alternative growth and leaching phases (Mode 2), spent medium leaching with separated growth and leaching phases (Mode 3), leaching with fresh growth medium without fungal culture (Mode 4), leaching with deionized water (Mode 5) and leaching with commercial organic acids (Mode 6). The main lixiviants produced in leaching Mode 1 were citric, tartaric, oxalic and gluconic acids in concentrations of 12, 2.5, 1.8 and 152 mM, whereas in leaching Mode 2, the concentrations were 15, 0.5, 1.0 and 1,162 mM, respectively. In leaching Mode 3, 63 mM citric acid, 23 mM tartaric acid and 29 mM oxalic acid was observed. At 5% pulp density, both leaching Modes 1 and 2 showed similar metal extraction yields from the electronic scrap at 35º C. Approximately 96% Al, 98% Zn, 48% Cu, 25% Fe, 66% Pb and 73% Ni were extracted in leaching Mode 1 in 14 days. In leaching Mode 3, 85% Al, 81% Zn, 97% Cu, 22% Fe, 50% Pb and 63% Ni were extracted in 14 days. The extraction yield of Cu was much lower in leaching Modes 1 and 2 (48- 50%) than in leaching Mode 3 (97%), as well as in leaching Mode 6 (98%) with organic acids, including citric, tartaric, oxalic and gluconic acids.IntroductionTechnical innovations in electrical and electronic equipment with new functionalities and designs stimulate consumer purchasing and lead a rapid increase in the production of waste materials. Terazono et al. (2006) have reviewed the data on electronic waste generation in different parts of the world. Studies by Bertram et al. (2002) have also affirmed that waste electrical and electronic components are the fastest growing waste category, which emphasizes the need for efficient recycling strategies. The recycling of electronic waste is an important subject, not only from the perspective of waste treatment, but also from the perspective of the recovery of valuable materials. However, the recycling of electronic scrap is still limited, due to the heterogeneity of the materials present in the products (Veit et al., 2005).Bioleaching-based processes offer a number of advantages over conventional pyro or hydrometallurgical methods (Lee et al. 2007; Lee and Pandey, 2012). These advantages include simple, inexpensive (in some cases) and ecofriendly approaches and the generation of less active leach residue relative to other metallurgical processes. Although biohydrometallurgical processes have been successfully applied to the leaching of metals from ores (Olson et al., 2003; Rawlings, 2002), data pertaining to the application of these processes for the extraction of electronic scrap are still scarce. Recently, a few studies have been undertaken for the extraction of metals from electronic scrap/ printed circuit boards (Brandl et al., 2001; Faramarzi et al., 2004; Ilyas et al., 2007; Ilyas et al., 2010). These studies were conducted with mesophilic chemolithotrophic bacteria (Acidithiobacillus ferooxidans and Acidithiobacillus thiooxidans), cyanogenic bacteria (Chromobacterium violaceum) or moderate thermophilic bacteria (Sulfobacillus thermosulfidooxidans). However, compared to bacterial leaching, fungal leaching has several advantages, including (i) the ability to grow at a higher pH, thus making it more suitable for the bioleaching of alkaline solid waste; (ii) a generally faster leaching process with a shorter lag phase and (iii) the chelation of excreted metabolites (e.g., organic acids) with metal ions, thus reducing toxicity to the biomass (Burgstaller and Schinner, 1993; Aung and Ting, 2005)."

Jan 1, 2015

By Robin Dill, Wei Song, Andrew Perham, Leo Martin

"AECOM completed a subsurface exploration program for an 18-foot diameter, approximately 4-mile long rock tunnel project in Hartford, CT. The tunnel crosses the regional geologic trend through the middle of a sedimentary basin, passing through four separate rock formations, including two volcanic flows. Faulting is prevalent, with at least 15 faults encountered during the site investigation. Severe artesian flows were also experienced in several of the borings, requiring highly specialized drilling equipment. Sixty-two borings and associated testing were performed to characterize the structural geology, hydrogeology, rock properties and ground conditions. A phased program was carefully planned and implemented with participation of geotechnical and geologic specialists, with thoughtful evaluation along the way to react to the ground conditions encountered, and make appropriate refinements. This paper will discuss the subsurface program details; the integrated efforts used to effectively characterize the ground; and will summarize the final characterization presented in the GBR. PROJECT DESCRIPTION The South Tunnel Conveyance and Storage Tunnel project is part of the Metropolitan District (MDC’s) Long Term Control Plan to address discharges to various water bodies throughout the greater Hartford area. This program includes a main conveyance and storage tunnel; a tunnel boring machine (TBM) launch shaft and retrieval shaft; adit tunnels with associated drop shafts and connection structures; consolidation conduits and manholes; and a tunnel pump station (TPS) at the downstream end of the main conveyance tunnel. AECOM is serving as prime designer for the MDC on this project. The main conveyance tunnel (hereinafter referred to as the “South Tunnel”) will be a deep rock tunnel with an invert depth of approximately 180 feet at its downstream end and approximately 218 feet at the TBM retrieval shaft. The tunnel will have a finished diameter of 18 feet. The tunnel will rise in grade at a slope of 0.1 percent to its upstream end. Total invert depth along the deep tunnel alignment ranges from about 180 ft. to as high as 320 feet. The total length of the tunnel will be 21,754 feet. The South Tunnel will be constructed using a shielded TBM and fully lined with pre-cast concrete segments."

Jan 1, 2016

By R. Zhang, S. Azam, S. Liu, J. H. Golbeck, V. Kurashov, S. Bhattacharyya

Free radicals (FRs) present in coal dust directly influence the interaction behavior between lung cells and dust particles. FRs are atoms or groups containing unpaired electrons, with strong chemical reactivity and a short lifespan. FRs which are quite persistent in natural environments, termed environmentally persistent free radicals (EPFRs), have recently received much attention as new environmentally hazardous substances. EPFRs present on coal dust serves as an active intermediary to catalyze reactive oxygen species (ROS) production by oxygen molecules endangering mine workers’ health. Many EPFR species may be bonded to the particle surface or trapped within the particle during coalification and can be modified with progressive size reduction through cutting at the mining face or crushing/pulverization in the plant. This study measured EPFRs for lignite, sub-bituminous, and anthracite coal dust at various sizes. We especially focused on EPFRs for submicron coal dust prepared by cryogenic methods. The outcome of this study will directly improve the free radical-dependent toxicity quantification and ultimately contribute to the size-dependent coal dust toxicity quantification.

Mar 2, 2022

By Peter Zhang, Daniel Su, Mark Van Dyke, Heather Dougherty

The integrity of shale gas wells in longwall chain pillars under the influence of longwall mining is a safety concern for both coal and gas operators. Previous modeling studies of gas wells in longwall pillars have shown that a FLAC3D model can predict longwall-induced stresses and deformations in gas well casings with reasonable accuracy. The predicted casing deformations can be verified by multi-finger caliper survey. This study concerns the influence of longwall mining on eight shale gas wells in a shale gas well pad located in the chain pillars between two adjacent longwall panels under 1,030 feet of cover. FLAC3D modeling was performed to predict longwall-induced deformations in gas well casings based on site-specific mining and geological conditions. Small deformations less than 0.5 in were predicted in the intermediate casing above 600 ft depth, and very small casing deformations less than 0.15 in were predicted below 600 ft due to high abutment pressure and high frictional resistance along the weak bedding interfaces. After the first panel was mined, a 56-arm caliper survey was performed on one of the wells to measure the longwall-induced casing deformations in the intermediate casing. The predicted casing deformations were then compared with the measured deformations, which indicated that the prediction was generally in good agreement with the measurement in terms of deformation locations and magnitudes. Findings from this study provide valuable information not only for stability assessment of shale gas wells in longwall chain pillars but also for developing engineering guidelines to preserve casing integrity through adequate pillar and casing design.

By Pascal Deschamps

"INTRODUCTION During the last decades, many underground facilities for the storage of hydrocarbons have been excavated worldwide and the technology is fully mature. Similarly to other underground structures, the initial basic design has to be adapted along with the work progress, in order to cope with the actual site conditions and in particular the geology, hydrogeology and associated rockmass quality. In addition, Client’s or underground Contractor’s constraints (such as resources availability, schedule or cost impact) may vary from design to construction and require last minute adjustments. Engineers are required to introduce flexibility and anticipation of unexpected situations (similar to the above-mentioned) while conceiving large underground facilities. Although most of the cases require only minor adaptations, crucial parts of the facility have sometimes to be considerably modified and reviewed. This article will present four cases evidencing how the design of one of these essential structures - the concrete plug – has been adjusted during site implementation without impairing the initial performances. Hydrodynamic containment The product containment of the unlined caverns discussed here is ensured by the principle of hydrodynamic containment. A permanent water flow is naturally or artificially maintained with an orientation toward the storage cavern and then, neither product-tight membrane nor lining are required. The hydraulic gradient (and the associated water flow lines), being directed from the surrounding mass toward the cavern, prevent from outward migration of the product. If the hydraulic gradient is not sufficient or if the water table is evidencing instability, water can be artificially injected into the rockmass through purpose-built boreholes, drilled from dedicated galleries. This water injection system, called the water curtain, is generally located several tens of meters above the level of the cavern. This principle is illustrated here after."

Jan 1, 2016

Water contact angles in the air/water/francolite, fuel oil/water/francolite and fuel oil/water/quartz systems were measured to describe the wetting characteristics of these minerals. It was found that white francolite is more hydrophilic than black and tan francolite both in the absence and presence of fatty acid addition. The contact angle measurement results do not correlate very well with the flotation response and it appears that kinetic aspects of the bubble attachment process play an important role in the francolite-fatty acid flotation system.

Jan 1, 1997

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Shrinkage stope mining can be employed in steeply dipping veins where both the wall and the vein material are sufficiently strong to stand with only minimal ground support. Shrinkage stope mining takes advantage of the swell in volume produced in breaking in-place ore. This expanded volume of ore is allowed to occupy part of the mined-out vein, and in so doing provides ground sup¬port and a working surface for the miners. Only enough of the ore is drawn after each stoping cut to allow work¬ing room. Not until the entire mining block has been broken is the full amount of broken ore drawn from the stope. DEVELOPMENT PRACTICES FOR SHRINKAGE STOPING The vein is developed into stoping blocks, each of which is 76 to 91 m (250 to 300 ft) in length and 61 m (200 ft) between levels. Level development, described later, is either on the vein or parallel to it with crosscuts periodically driven over to the vein. There are usually two exits from each stope. The first entry is a raise at one end driven either conventionally by raise climber, or by a raised borehole. Present practice is that this raise, put in from level to level, is either a 2.1- x 2.1-m (7- x 7-ft) Alimak raise or a 1.2- to 1.8-m (4- to 6-ft) diam borehole. Either raise system will require pro¬visions for an ore chute carried with the stope to the haulage level or the scram drift. At 9- to 12-m (30- to 40-ft) vertical intervals, dogholes are developed from the raise so that when stoping breaks back into these dogholes, timber or steel sets and necessary service facilities can be installed. The second entry to the stope may be either a similar raise at the stope endline, or more economically, a cribbed raise carried with the stope. As previously stated, development of the ore block is achieved by either drifting on the vein or by paralleling the vein with a footwall lateral and then crosscutting over to the vein at specified intervals. In the following discussion, three basic cases are illustrated in Figs. 1, 2, and 3. In Fig. 1, development is in the footwall by a 3- x 3-m (10- x 10-ft) trackless lateral. A heading is also driven on the ore and taken either full width or driven a standard size and slabbed to the vein width. Approxi¬mately 9-m (30-ft) long crosscuts on 7.6-m (25-ft) cen¬ters are driven from the lateral to the vein. For ease in muck handling they are usually put in on a slant rather than at right angles. A single orepass, located off the lateral and in the center of one stope block, can serve not only that stope, but also one on either side, up to a tramming distance of not more than 137 m (450 ft). Raises are put at the stope endlines on 91-m (300-ft) centers. The back of the vein is shot down and just enough broken material is drawn off to permit access for men to work off the broken ore. Development in Fig. 2 progresses by driving a 3- x 3-m (10- x 10-ft) trackless heading on the footwall side of the vein. The ore block is delineated by raises on 91-m (300-ft) centers. Crosscuts are again driven on 7.6-m (25-ft) centers over to the vein. Approximately four rounds are then taken up to the vein and each round is successively belled out so that all the crosscuts (pockets, in this case) connect. In this development, as opposed to that in Fig. 1 where there were no belled

Jan 1, 1982

By D. A. King

Monitoring devices have been installed at several mine sites to assess the potential for the development of acid rock drainage (ARD) in waste rock, spent ore, and tailings material. In addition, instrumentation has been designed for monitoring changes in water quality in pit lakes. These devices monitor temperature, oxygen, water movement, and water chemistry in the waste rock pile. This information can be used to evaluate the current and future ARD risk, the effectiveness of caps and covers, and provide information to aid in the design of other reclamation strategies that limit ARD development. The instrumentation includes nested thermistors and oxygen sensing ports, neutron access tubes, heat dissipation units (HDUs), and lysimeters. Many of the monitoring devices can be linked to a datalogger and telemetry systems, enabling monitoring data to be downloaded from the office.

Jan 1, 1997

By M. R. Shahan, W. R. Reed, J. P. Rider

Roof bolter operators may be exposed to high respirable dust concentrations on continuous miner sections with blowing face ventilation when bolting is performed downwind of the continuous miner. One solution to reduce the high respirable dust concentrations is to use a canopy air curtain (CAC) to deliver clean air from a filtered blower fan directly to the bolter operators under the canopies. The influence of CAC installation in the airflow and dust dispersion around the location of the roof bolter operator can be evaluated by using computational fluid dynamics (CFD). This study, performed by the National Institute for Occupational Safety and Health (NIOSH), considers two scenarios: (1) a roof bolting machine in the center of the entry for installation of the fifth row of bolts from the face, and (2) a roof bolting machine positioned close to the face for the installation of the last row of bolts. In both scenarios, the bolting machine is placed in an environment which contains 6.0 mg/m3 of respirable dust and is ventilated by a blowing curtain with 3000 cfm (1.42 m3 /s) of air. This environment is used to simulate the roof bolter machine operating downstream of a continuous mining machine. Two operation positions are simulated at the same bolting location: dual drill heads in the inward position for two inside bolts and dual drill heads in the outward position for two outside bolts. The influence of the CAC on airflows and dust dispersion is evaluated with the CAC operating at 250 cfm (0.12 m3 /s).

The latest resource estimate for the Caspiche gold/silver/copper project in the Maricunga region of Chile was released by gold explorer Exeter Resource Corp. The resource is low grade, but bulk minable and drilling to date has shown a NI 43-101 compliant inferred resource of 1,117 Mt (1.23 billion st) grading 0.55 g/t (0.016 oz/st) gold, 3.81 g/t (0.111 oz/st) silver and 0.22 percent copper. Exeter said these findings make the in situ metals amounts of 609 t (19.6 million oz) of gold, 4.4 kt (137 million oz) of silver and 4.84 billion lbs of copper, which the company relates as 33.7 million gold equivalent ounces using assumed metal prices of $23.33/g ($800/oz) for gold, 35 cents/g ($12/oz) for silver and $2/lb for copper. The deposit is almost midway between Barrick?s Cerro Casale gold deposit and Kinross? Refugio Mine. In a statement Exeter chairman Yale Simpson said ?This updated resource estimate places Caspiche firmly among the world?s largest gold-copper discoveries in recent years and clearly supports our decision to aggressively drill the property, despite the global economic decline. ?We expect that the very substantial copper and silver credits will be extremely important, and will markedly impact the potential viability of Caspiche. Specifically, using metal prices of US$2/lb copper, $23.33/g ($800/oz) gold and 35 cents/g ($12/oz) silver, the 4.84 billion lbs of copper and 4.2 kt (137 million oz) of silver are equivalent to 376 kt and 62 t (12.1 and 2 million oz) of gold, respectively. ?Our recent drilling has defined what appears to be a coherent higher grade central zone associated with a definable early stage diorite intrusion. We expect this zone will be important in developing various mining and economic models for the project,? Simpson said. Exeter is still drilling the deposit to expand the re-source estimate further to the southeast and at depth and to upgrade the higher grade central zone to the indicated category. Simpson said, ?Exeter intends to enhance the value of Caspiche by continuing to de-risk the project through ongoing exploration, metallurgy, engineering, water and environmental studies. These studies will continue through the next 12 months and will lead to a conceptual development study. A timeline for the release of that study is being formulated. The project budget for the next 12 months is C$14 million.? Mineweb reported that Caspiche is likely too big of a deposit and at too high of altitude for Exeter to handle alone. Exeter will be looking to other majors, or perhaps the Chinese, to help develop this huge orebody. But, in the meantime, it also has its high grade Cerro Moro deposit in Argentina, which could be brought to production relatively quickly and at relatively low cost.

Jan 1, 2009

By J. W. Jang, I. Iwasaki

Introduction Grinding ball wear in wet grinding is due to a complex combination of abrasion, corrosion, impact and their interactions. The ball charge is continually falling from a height greater than half of the ball mill's diameter, and contact is made with ore particles or other balls. Abrasive wear occurs through penetration and plowing of media materials from the ball surface by hard ore particles (Gangopadhyay and Moore, 1985). Electrochemical interactions between the abraded and unabraded areas of the ball surfaces, as well as between the grinding media and the minerals, occur in grinding mill pulps (Adam et al., 1986). The wear depends on the media composition, the hardness and the corrosive/abrasive characteristics of the slurry. The marked-ball-wear test method was used to evaluate the performance of grinding media under nitrogen and oxygen atmospheres. Also, the wear mode was studied by observing the grinding media surface under a scanning electron microscope. In general, grinding media of high hardness and high chromium contents were found to have better corrosive wear resistance than those of low hardness and low chromium contents (Jang, Iwasaki and Moore, 1988). The purpose of this investigation was to determine the size distribution and the shape of wear debris in a study of the wear rates and wear behaviors of grinding balls. Experimental Materials The grinding ball alloys used in this investigation were: •martensitic stainless steel (MSS; 0.96% C and 16.2% Cr), •high-carbon, low-alloy steel (HCLA; 0.89% C and 0.53% Cr), and •mild steel (MS; 0.22% C and 0.09% Cr). The mineral used in this study was synthetic corundum, because it was on hand and contained no magnetic impurities. Debris A 203- mm (8-in.) diam porcelain mill was used for grinding the corundum. The mill charge consisted of: •1200 g of 12-grit synthetic corundum, •770 mL of 0.05 mole/L Na2SO4 (61% solids), and •125 steel balls of nominally 25-mm (1-in.) diam. To minimize the debris corrosion, nitrogen was introduced into the mill. Grinding with MSS, HCLA-steel or MS balls was conducted for 2 hr at 51 rpm. The grinding time using the MS ball was varied from 0.5 hr to 2 hr in an attempt to study the debris amount and shape as a function of time. The debris was collected with a hand magnet from the ground slurry and separated from the corundum particles using a Davis magnetic tube. After the separation, the debris particles were immediately washed with acetone and then filtered and dried. Debris size analysis Buckbee-Mears 3-in.-diam, micro-mesh sieves and an electromagnetic shaker were used for determining the debris size distribution. In sieving, isopropyl alcohol was used. The debris fraction collected from each sieve was washed with acetone, dried and then weighed on an analytical balance. [ ]

Jan 1, 1993

By Y. Kurokawa, R. Kurosawa, I. Iwata, I. Fukuoka, N. Sugita

INTRODUCTION The Ningyo-toge Mine is located in the western part of Honshu, about 800 km from Tokyo. Since the discovery of uranium ore deposit at Ningyo-toge in 1955, different kinds of mining and milling tests have been carried out. A new pilot plant of milling and UF6 conversion is also being under construction. Operations are scheduled to start by the end of this year. In 1979, the uranium enrichment pilot plant was constructed at Ningyo-toge and test operations have started. Thus, the Ningyo-toge area is assigned to undertake the Research and Development of the upstream of the nuclear fuel cycle in Japan. PNC has been strengthening its security measures more than ever. This paper will focus on the disposal of wasted material, especially on the disposal of heap leaching residue, in Japan. URANIUM ORE DEPOSITS The uranium of the Ningyo-toge Mine is characterized genetically by epigenetic sedimentary type deposits. They occur in the host rock of Neogene tertiary fluviatile conglomerates and sandstones which unconformably overlies the Mesozoic-Palaeogene granitic basement covered with lucustrine deposits and volcanic ejecta. More than 12 deposits have been found over the area of approximately 160 km2, and distribution of the deposits are strongly controlled by the so-called channel structure, probably remnants of paleo-streams on the denudation surface of the underlying grantic basement. Each ore body has different shape and dimension respectively, but about 2-3 meters thickness with the average grade of 0.05-0.2% U308 is common feature of the ore bodies. The ore is soft and loose, and it is composed of common rock forming minerals of allogenic origin and their altered products. The chemical composition of the ore depends on the constituents of the country rocks or the pebbles. A primary mineral, ningyoite and secondary mineral, autunite are major ore minerals, although three primary tetravalent minerals and thirteen secondary hexavalent minerals have also been detected. MINING EXPERIMENT Since 1959, various types of underground mining tests have been carried out at Ningyo-toge Mine. The room and pillar method was safe and simple, although the uranium ore recovery was below 70%. After experiment of the retreating long-wall method and slicing mining method, it was found that the modified long wall method was the most appropriate for the underground mining of the Ningyo-toge uranium deposits. This method was completed successfully to increase efficiency up to 2 or 3 times, as compared to conventional methods. MILLING EXPERIMENT In 1964, a small test mill was constructed at Ningyo-toge, and a milling experiment using the Ningyo-toge uranium ore was started. Re-modelling of equipments and improvement of the process lead to the establishment of the technique of the so-called wet process, and produce UF4 directly from the ore. The flow diagram is shown in Fig. 1.

Jan 1, 1981

By M. G. Eiselein, R. A. Smith, C. H. Sheerin

FMC Gold's Paradise Peak Mine was one of the top two producers of mercury metal in North America in 1988 and 1989. As a by-product, elemental mercury is produced from cyanide soluble mercury minerals, primarily to reduce the health risks incurred with the refining of precious metal precipitate containing elemental mercury and mercury amalgams. A discussion of the design, primarily the Merrill-Crowe units, and permitting for the safe production of mercury will be presented. Additionally, laboratory testwork to determine the effect of mercury on reagent requirements in the Merrill-Crowe process will be discussed.

Jan 1, 1990

By Scott F. Andersen

Discussion to include location and description of the project from upper lake through powerhouse to lower lake, noting locations and size of shafts. Discuss evolution of shaft development concepts in conjunction with owner, contractor and CAL OSHA. Discuss elevator shaft equipment, T1 gate shaft, T2 surge shaft equipment, T3 surge shaft and the incline shaft equipment. Question and answer period.

Jan 1, 1981

By S. Pollak, J. Rickard

The Long Baseline Neutrino Facility Far Site in Lead, South Dakota is currently under construction and involves drill and blast excavation of the largest and deepest caverns in North America. This paper will review how the geotechnical risks are being managed through careful consideration of excavation sequence, completion of exploratory pilot tunnels, implementation of a robust instrumentation and monitoring program, and validation against the design. Aside from having a cross section of 5,900 square feet and being situated at a depth of one mile, the caverns also employ a unique permanent rock bolt solution which underwent a significant pre-production pull testing program involving varying installation methods and being subject to blast vibrations. The behavior of the schistose rock mass and performance of the ground support are being continually monitored throughout the construction phase.

Jun 13, 2023

By Terry D. Vandell

The purposes of this presentation are: (1) to discuss the actual as well as potential incidents that have caused and that could cause groundwater contamination at uranium in situ leach mining sites and (2) to discuss remedial actions and solutions that have been or that could be implemented in the future to stop or prevent serious groundwater contamination from in situ leach mining in Wyoming?s Powder River Basin. Open pit, underground, and existing and proposed in situ leach uranium mine sites in the Powder River Basin are shown in Figure 1. In situ leach mining or solution mining for extracting minerals in liquid form is a relatively new addition to the list of conventional mining methods currently used to extract uranium in Wyoming's Powder River Basin. Basically, the in situ leach mining method involves: (1) the injection of a leach solution lixiviant (an acidic or basic oxidizing solution) into a uranium-bearing ore body, via injection wells, to complex the contained uranium, (2) mobilization of the uranium complex from the host material via creation of a soluble complex salt, and (3) surface recovery of the solution bearing the uranium complex via production wells. Uranium is then separated from the leach solution by conventional milling unit operations (ion exchange).

Jan 1, 1981

By G. M. Swinkels

It is generally accepted, that oxygen is a necessary reagent for the flotation of galena with xanthate. Recent measurements have shown (1) that under plant conditions oxygen may be in a short supply to the extent that it limits the rate of flotation. Thus, it becomes important to define the exact role of oxygen and establish a model for the rational design and control of flotation separations, involving oxygen as a reagent. All theories or, the flotation of galena assign oxygen the role of an oxidant. However, they disagree on a number of questions: 1) Is the galena or the xanthate being oxidized? 2) Is a lead dixanthate or dixanthogen the actual collector? 3) Is the oxidation process in which oxygen participates of a chemical or an electrochemical nature? The last question arises from the observation (2) that an anodic bias as well as oxygen will make a galena surface hydrophobic in the presence of xanthate. Theories which are based an the premise that the influence of oxygen is of an electrochemical nature offer elegant models of galena flotation. However, full advantage has not been taken of available electrochemical techniques to discriminate between theories and this is the area towards which this research has been directed (3).

Jan 1, 1970

By William R. Taylor

The information I have concerning the early in situ uranium mining activities in Texas is "sketchy." Apparently, the first test was conducted between 1967 and 1971 in Gonzales County, north of Yorktown, at the Dueterstat Site, by, City Service. Other tests were run in Webb County at the Bruni Site, and in Live Oak County at the Burns and Clay-West Sites in the early 1970's. During this period the Texas Water Quality Board maintained a general surveillance of these tests. In mid-1974, Mr. Glen Davis walked into our Austin offices and requested a permit for Atlantic Richfield's first in situ uranium mine. This permit, issued in January, 1975, was for 3 acres of production, and was the first such permit issued in the United States. It formalized the regulation of in-situ uranium mining in Texas. The 4 years since then have been characterized by growth and change. Today there are 20 permits approved for approximately 2,000 production acres. These permits are issued to 8 companies at 13 sites in 6 counties. There are 7 additional permits being processed, or on the verge of being applied for. These will involve possibly another 2,000 production acres, 3 more companies in one more county.

Jan 1, 1979

By J. J. Kellar

Researchers have long speculated regarding the adsorption state of collectors on mineral surfaces. Only recently have analytical techniques become available which allow direct in-situ analysis of adsorbed layers. Fourier transform infrared spectroscopy coupled with internal reflection spectroscopy (FT-IRJIRS) with reactive internal reflection elements allows for unobtrusive investigation of the adsorbed layer (Harrick, 1979). This technique has been used both for the in-situ identification of adsorbed species and for the direct calculation of collector adsorption density in real time from infrared spectral data (Miller and Kellar, 1988). The reactive mineral IRE serves both to internally reflect the infrared light and to act as a substrate at which collector adsorption can be monitored. Shown in Figure 1 is a schematic of light being internally reflected in an IRE. As the light is totally reflected at the interface, an evanescent wave propagates a short distance (typically a few microns) away from the IRE, and sampling of the interfacial region occurs. Because of multiple internal reflections, great sensitivity can be obtained. Flotation systems for which this method has been demonstrated include CaF2/oleate, Al2O3/ dodecyl sulfate, CaCO3/oleate, ZnS/xanthate, and KCl/octylamine (Miller and Kellar, 1988; Kellar et al., 1989, 1990; Miller et al., 1990).

Jan 1, 1992