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By Guangming Zhao, Nianjie Ma, Demao Guo, Denghong Chen, Yingming Li

Based on the mechanical essence that large-scale plastic failure zone appears in all or part of surrounding rock in soft rock roadway, the numerical simulation method is used to study the rectangular roadway in layered rock strata. It is clarified necessary conditions must be met for soft rock: firstly, the strength condition is that the maximum confining pressure is greater than the uniaxial compressive strength of rock strata. Secondly, the stress environment condition is that the ratio of maximum confining pressure to minimum confining pressure is greater than 3. Thirdly, the angle condition is that The direction of principal stress action enables the plastic zone of weak rock layers to fully develop. At the same time, the quantitative description method of soft rock is given, and the soft rock roadway is redefined. Soft rock roadway refers to the roadway that meets the strength conditions, stress environment conditions, and rock structure angle conditions under certain surrounding rock conditions and in-situ stress environment conditions. After the excavation of the roadway, a large-scale plastic failure can be formed, that is, a butterfly-shaped plastic zone is formed, and the conventional support cannot be adapted. It is difficult to support in engineering. It provides a theoretical basis and engineering analysis method for the identification of soft rock roadway, and the research results have engineering value Soft rock tunnel engineering in coal mines constitutes a vital aspect of soft rock engineering. This field broadly encompasses rock engi- neering concerning large plastic deformations, e.g., soft rock slope engineering and soft rock tunnel engineering. The intricate geological conditions encountered in soft rock tunnel engineering present a significant challenge to support, which has harmed coal production in China. China leads global raw coal production with the annual output of 4.6 billion tons. Annual tunnel excavation supporting this production spans approximately 11,000 km, with over 10% of these tunnels classified as soft rock formations. Soft rock is commonly associated with soft rock tunnels due to their prevalence in engineering projects. However, reaching a consensus on the definition of soft rock has long been an enduring challenge for scholars and engineers. Numerous definitions have been proposed, includ- ing descriptive, index, and engineering definitions. For instance, the International Society for Rock Mechanics defines soft rock based on its uniaxial compressive strength σ ranging from 0.5 to 25 MPa. China's Engineering Rock Body Standards, established in 1994 (GB 50218-94), take a qualitative and quantitative approach to classifying rocks. Rocks are categorized as hard or soft based on criteria such as hammering sound, fragmentation, water immersion effects, and weath- ering degree. Additionally, the integrity of rock bodies is assessed across five categories intact, relatively intact, soft fractured, fractured, and extremely fractured. This classification considers factors like the number and spacing of structural planes, their combination, and the types of structures. Descriptive and index-based definitions fall under the category of geological soft rocks, providing a comprehensive geological perspective on the surface features or strength characteristics. However, these definitions have limitations in engineering practice, which leads to contradic- tions. For instance, rocks with uniaxial compressive strength less than 25 MPa may not exhibit soft rock characteristics if the tunnel is shal- low with low horizontal stress levels. Conversely, rocks with compressive strength exceeding 25 MPa at sufficient depth and high horizontal stress may exhibit soft rock characteristics. Definitions originating from engineering practice have emerged after realizing the inadequacy of discussing soft rocks without considering engineering. For instance, Dong's loose circle theory defines soft rocks as rocks with a loose circle thickness exceeding 1.5 m, which chal- lenges conventional supports. This intuitive definition, widely accepted by engineering professionals, emphasizes the difficulty in supporting tunnels due to extensive damage. However, various tunnel damage poses a challenge in relying solely on the loose circle thickness of tunnels for determining soft rocks. He introduced the concept of engineering soft rocks, which are defined as rock formations exhibiting significant plastic deformations under applied engineering force. Two fundamental mechanical properties of soft rocks are identified the critical softening load and critical soft- ening depth. Rocks below the critical softening load threshold are categorized as hard rocks, while those exceeding it exhibit substantial

Jun 25, 2024

By Bob Snashall

"We could blow that damned fort out of existence if we could run a mine shaft under it!," claimed one Union soldier eyeing Bobby Lee's Petersburg defense line protecting Richmond and Jeff Davis. Puncture the line, take the hill behind, and the Confederate capital would celebrate Yankee jubilee, 1864. That was June. At 4:15 am, July 30, mine boss Henry Reese, sergeant, 48th Regiment, Pennsylvania Volunteers, gingerly entered the mine, dark and muddy. Something had gone wrong. The mine, set to go off at 3:30, hadn't exploded. It had to be checked. The main mine gallery, 1.5 m high x 1.4 m wide (5 x 4 ft) narrowing at the top to 0.6 m (2 ft), forced Reese into a tuck run. Some¬where up there in that 155 m (511 ft) stretch running 6 m (20 ft) under the rebel works, moisture may have doused the charge. Or maybe a fuse splice went awry. Or maybe there was live fire steaming like a locomo¬tive towards 3.6 t (4 st) of black powder. Perhaps now Reese wished this sprint was shorter. For the mine boasted a record break¬ing distance in military annals. Lieut. Col. Henry Pleasants Credit goes to Lieut. Col. Henry Pleas¬ants, commander of the 48th, who overheard the soldier's claim for what a mine could do. In peacetime, Pleasants had been a railroad engineer working on the Sand Patch Tunnel. He later switched to mining, working for the Philadelphia & Reading Coal & Iron Co. He even married the daughter of the editor of The Miners Journal. Grieving over his wife's early death, Pleasants plunged into the ven¬ture of war. This Union officer took the mine idea to the brass. Headquarters was less than enthu¬siastic. "The chief engineer of the Army and the rest of the regular Army wiseacres said it was not feasible," Pleasants wrote to his uncle. The pros scoffed, ticking off a list of horribles: No mine has ever gone that far, men will suffocate and be crushed, the rebs will countermine, the men will be trapped. Still, Pleasants eked out a signal to proceed. In the meantime, the miners of the 48th had commenced to dig anyway. It was like the 48th to proceed against the odds. The 48th Pennsylvania Volunteers The 48th had been recruited from Penn¬sylvania's Schuylkill County, anthracite coal mining country. The mines had been torn with strife over wages. Strikes and rioting were prevalent. Murder was not unknown. This "Buckshot" resistance centered in Pottsville, PA. It forced the governor to call in troops. Coal was in high demand to run the war machine. Pennsylvania operators worried that a military draft would stir up violence, closing the mines altogether and further shift¬ing business to emerging Illinois mines. Men lit out for the western gold fields in the tens of thousands. Secret societies bred in the mining camps invited draft agents to dance at the end of a rope. Jobs threatened by cheap, victorious South¬ern labor were one warning intended to get immigrant miners interested in the war. Operators gave miners time off to help de¬fend Philadelphia during Lee's thrust to Gettysburg. Even so, the draft spawned new violence and troop intercession. Soldiers were not welcome. The dust had barely settled when into this Pennsylvania "minefield" marched the 48th on a recruitment drive. The 48th had been in Lexington, KY on guard duty. When it came time for the regiment to pull out, the townspeople had prevailed upon authorities to extend the regiment's stay. Finally ordered to move out, the 48th was feted by a hanky-waving, tear-dropping send-off to the lyrical lament Auld Lang Syne. And Kentucky fancied itself to be rebel. Striding smartly into the Pennsylvania coalfields, the 48th attracted 400 recruits and marched off under regimental colors specially made by the proud citizens of none other than Pottsville, PA.

Jan 1, 1989

By Peter V. Avotins

During last year's AIME meeting in New Orleans, Deepak Malhotra and I were discussing some of the topics that would be presented in this Symposium. It seemed appropriate that along with papers on specific reagent and flocculant types, we should address research management aspects of new product development. This development is a long, arduous task that requires a significant technical and fiscal commitment by the chemical company and potential users. A flow chart depicting the research process is shown in Figure 1. The sequence of events are plotted on an undefined time line. Our experience has shown that it takes 3-5 years to complete a typical reagent development project. A new project starts with an idea of how to do something better. The ideas range from relatively simple product modifications to opportunities created by technological change. An example of this is the use of flocculants in tailings operations to close the water recycle loop. The change was brought about by the need to conserve water and comply with environmental statutes. The result was the development and wide-spread use of synthetic polymers to help in solid-liquid separation. More recently the ups and downs of oil prices have led to development of coal water slurries as an alternative to fuel oil. The slurries contain high levels of chemicals (up to 3%) to insure proper stability and burn characteristics. There are technical changes under way right now that may lead to ideas for other new products. One exciting source of new technology is the ability to manipulate genetic material of microbes to select for specific capabilities. This has evolved by leaps and bounds over the past 5 years. Biologically assisted leaching and removal of wastewater impurities are two possible end uses. I'm sure that there are many others. The ideas can originate anywhere and are often half-baked at first, but to lead to new products they must be refined to provide project definition and eventually a research plan. A chemical ,development project cannot be started without approval to commit the company's resources. The internal selling job is often very difficult. The company must see clear financial benefits in order to launch efforts that may not pay off for many years. It is critical to project the total market, profitability, eventual market share, and timing of introduction of the new product. There is much uncertainty in this and the research manager is often skating on thin ice to retain his credibility. Once we obtain support from upper management, the research team is assembled to start the project. The teams generally consist of organic chemists, research metallurgists, and technicians. They have access to analytical scientists and technicians for surface and analytical chemistry. The projects are often augmented by help from university staff and consultants. This is particularly true where there is a need to develop fundamental understanding of the technology in addition to synthesis of chemicals. The project now proceeds along two lines. First, there is an emphasis on understanding the prior art. Extensive literature and patent searches are undertaken to provide research leads and to keep from "reinventing the wheel" and infringing on someone else's patent. The scientists and technologists are qualified to conduct computer-assisted library searches. This process usually uncovers 80-90% of the relevant information. The remainder is very much harder to get and requires the help of professional library search staff with scientific backgrounds. Along with literature searching, the team begins synthesis of new compounds. These are screened by use of techniques such as microflotation or filtration in a Buchner funnel. The scientists are looking for the direction of response as they modify structural features . Inevitably, a large number of candidates reach the next stage, a laboratory bench test. At this point it is possible to set up a research plan and use project management techniques to monitor progress. The bench tests are conducted in larger equipment, usually requiring 100-1000 grams of test substrate. Typical laboratory flotation tests are run in Denver or Wemco cells.

Jan 1, 1986

By P. K. Hopke, A. Hubbard, K. H. Leong, J. J. Stukel, K. Nourmohammadi

INTRODUCTION In order to understand the transport and deposition of radon daughters in mine atmospheres, it is necessary to know the variation in the attachment of the daughter atoms to particles as a function of particle size, composition, number density, relative humidity, temperature, and radon concentration, the free gaseous diffusion coefficients of the daughters, and the variation in the mass transfer of the activity, both free and attached to particles, to mine surfaces as a function of particle size distribution, surface roughness of the mine walls, and the flow conditions. If all of these parameters are known in a model system, it should be possible to understand the transport and fate of the airborne radioactivity in real mines under certain well-defined flow conditions. There have been a number of recent investigations of the attachment of radon decay products to particles 1-4, but there are still a number of unanswered questions regarding the process. However, it is clear that for most real mine atmospheres, the vast majority of the activity is attached to particles. The size distributions for the activity-bearing airborne particles have been studied 5,6, and it has been found that most of the activity resides on particles with diameters in the range of 0.05 µm to 0.3 µm with an average mass median diameter between 0.1 and 0.2 µm. The behavior of the unattached radon daughter species has also been recently studied[ 7], and many of the previous problems regarding the value of the diffusion coefficient for Po-218 have been resolved. A major problem in the understanding of the airborne transport of radioactivity in mines is the lack of detailed knowledge of mass tranfer to and fluid flow over rough walls under fully developed turbulent flow conditions. This paper will report the progress on a project that is designed to obtained that information. MATHEMATICAL MODEL DEVELOPMENT Deposition of particles on smooth surfaces in turbulent flow has been extensively studied. A comprehensive review of these results has been prepared by Sehmel 8. There has not been such a comprehensive study of particle deposition on rough walls under such flow conditions. In recent years, only a single model has been proposed to explain such deposition 9,10 and in both of these papers the flow structure in the rough walled pipe was not taken fully into account. As part of the work being conducted on this project, a more complete model was outlined in a previous report [11]. The basic theory will be reviewed to provide a context for the flow measurements to be reported. The flux of particle deposited on the walls of a pipe in a turbulent flow is derived from the one dimensional form of Fick's law as given by [N = Dpdpp/dr (1) where N is the flux of particles deposited per unit area per unit time, D is the total eddy diffusivity of the particles, p is the airborne concentration of particles, and pr is the distance measured from the center of the pipe. The rate of deposition is best expressed by a deposition velocity Vp = NIP pb (2) where P b is the mean particle concentration in the sulk flow. The shear radius, V/ut and the shear velocity, u , are used to calculate a nondimensional distance, and velocity, respectively, where v is the kinematic viscosity of the fluid. The nondimensional form of equation 1 is given by Vd = DP dpp(3) V dr+ where Pp = Pp/ Ppb(4) By integrating equation 3 from the rough wall stopping distance, S , to the center of the pipe, the deposition velocity can be obtained. In order to make this calculation, it is necessary to have accurate descriptions for the particle eddy diffusivity, stopping distance, and shear velocity in order to insure that the influence of the flow structure has been properly accounted for. The shear velocity can be determined experimentally from the shear stress evaluated at the wall, Tw, and the fluid density, ut =VT w/p = ub V f/2 (5) where ub is the mean bulk axial velocity. The wall shear stress for a given pressure drop, dP/dL, and hydraulic diameter, Dh, is]

Jan 1, 1981

By W. J. Schlitt

Introduction A major uncertainty in assessing the potential for heap-and dump-leach projects is how to determine the extraction-rate curve for the recovery of the mineral values from coarse ore. Such material could either be run-of-mine (ROM) or primary crushed ore. The problem with field testing coarse ore, especially for new projects, is the large scale and extended leach times needed to accurately determine the final extraction-rate curve. At least 5 x 103 to 5 x 104 t of representative ROM ore are typically required for a copper test heap, and much more is often used. Kennecott, for example, recently constructed a 0.9 Mt (1 million st) ROM test heap at the Bingham Canyon Mine in Utah. In such coarse ore operations, the ultimate level of extraction will require a leach cycle that can extend from several months to a few years. Quite often, project development schedules do not provide the luxury of mining such large quantities of material or running such long tests. Instead, test data are usually limited to results from column leach studies on relatively fine ore, often with a top size that does not exceed 25 mm. Maximum leach times are also short, typically less than a year before an initial decision is needed on project viability. Proposed method One approach to estimating the recovery from a coarse ore leach is to assume that the leach solution will have some ultimate penetration distance into the rock. Then, the final level of mineral extraction in this "leached rim" will equal the ultimate level of extraction identified in various testing programs. Obviously, if the radius of a given rock fragment is less than the penetration distance, that fragment will be fully leached at the end of the operation. With larger rock sizes, the percent recovery will fall off as the size increases and the fraction of unpenetrated rock mass increases. Such an approach sounds simple but is likely to be complex when applied to a real project. For example, the penetration distance will be a function of both the rock characteristics and the effective length of the leach cycle. The important rock characteristics include rock porosity, the degree of internal fracturing and the mode of mineral occurrence. With regard to the latter, penetration is likely to be greater if the leachable mineralization occurs on fracture surfaces or in veinlets, as opposed to fine grains uniformly disseminated throughout the rock mass. An estimate of penetration distance may be derived from column or heap tests by noting the depth of solution penetration into the larger rock fragments after three, six and 12 months of leaching. While the penetration rate is ore specific, something on the order of 10 to 20 mm/y may be appropriate for competent, primary copper (chalcopyrite) ore. For gold in tight quartz, the rate may be about the same. Copper oxide ores and gold that is hosted in a more porous rock matrix are likely to have penetration rates that are at least two to three times higher, and an even higher rate should be appropriate for uranium hosted in sandstone. As noted above, the length of the effective leach cycle is likely to be measured in years. On this basis, the ultimate penetration distance (dp) would vary from less than 50 to several 100 mm when a particular ore is leached to exhaustion. Several sets of mathematical manipulations are necessary to convert a rock size distribution and corresponding value of dp into an estimated extraction-rate curve. The first step is to break the ROM size distribution down into intervals and then calculate the radius for the mean rock size in each interval. This is shown in Table 1 for rock sizes up to 1.75 m (about 6 ft) in diameter. The next step is to calculate the volume of unleached core and the fraction of rock that is leached. This is done for the following three values of dp: 25, 100 and 250 mm. Results are shown in Table 2. The third step is to select the ultimate level of recovery that will be achieved in the fraction of material that is effectively leached, i.e., the outer zone that is penetrated by the leach solution. This is clearly a site-specific factor that can only come from metallurgical test results on representative ore

Jan 1, 1998

By R. H. Moore, B. D. Breitenstein

INTRODUCTION The United States Uranium Registry (USUR) tissue program was funded in 1978 by the United States Department of Energy (DOE) through its Human Health and Assessments Division, and registrant enrollment began in 1980. It is operated by Hanford Environmental Health Foundation (HERF), Richland, Washington, with support provided by Pacific Northwest Laboratory (PNL).** The USUR tissue program is the first systematic effort to contact and gain assistance of workers occupationally exposed to uranium and to seek and study uranium deposition in human tissue. This paper describes the program objectives, populations selected for study, procedures for enrolling workers in the program, types of tissue to be studied, and analytical procedures to be used. PROGRAM OBJECTIVES The objectives of the USUR tissue program are to: 1) Determine the distribution and concentration of uranium, if any, in the tissues of occupationally exposed workers. 2) Compare bioassay measurements of exposed individuals with the results of analyses of tissues obtained at autopsy. 3) Seek evidence of histopathologic changes related to any uranium deposition that may be present. 4) Conduct analyses of whole bodies, when available, to obtain more precise data on the uranium burdens, if any, in the body and organs, and especially the distribution in parts of the body, such as the skeleton, that are not usually accessible for sampling. 5) Develop data that will assist in evaluating a) the accuracy of current in-vivo measurement techniques, b) the propriety of existing regulations, and c) the adequacy of current protection programs. SELECTION OF POPULATIONS FOR STUDY Selection of the populations for the USUR tissue program was based on studies of the United States Uranium Registry described elsewhere (0c81). Potential populations for study were quite varied because of occupational exposure to different chemical forms of uranium, different levels and eras of exposure, varying ability to identify populations exposed in the past, and the general interest and cooperativeness of the populations. Nevertheless, certain identifiable populations emerged that were both willing to participate in the study and likely to provide useful information. Among the targeted populations are individuals who were exposed to uranium prior to the existence of modern exposure standards. These individuals are presently being enrolled by the USUR. The reliability of concentration and distribution data and the demonstration of effects are more likely in such individuals than in those who have received less exposure. However, the targeted groups also include individuals who Have been exposed to permissible levels of uranium, and reference employees in uranium plants who have not worked with uranium. The lesser exposures are of immediate interest because of the rarity of heavy exposures under current standards. The sampling of potentially exposed individuals with negative in-vivo measurement results may serve to check the adequacy of in-vivo measurements in the detection of deposition. [ENROLLMENT PROCEDURE] The enrollment of participants follows three patterns. 1) Presently employed workers are enrolled by the cooperative efforts of USUR staff with the medical, industrial relations, and/or health physics personnel of the uranium facility. Volunteers may be enrolled by medical personnel at the time of their physical examination or by health physics personnel at the time of a bioassay or lung counting procedure. Follow-up after initial contacts with currently employed uranium workers is carried out by on-site personnel, but subsequent follow-up is the responsibility of the Registry. 2) Retired workers are contacted by mail, with a covering letter from their former employer supporting the principles of the study but leaving participation in the program on an entirely voluntary basis. 3) An alternate method of enrollment is to obtain a USUR autopsy consent after death from the next of kin. As part of enrollment in the program, permission is requested for access to the individual's medical and exposure data, and a short occupational history is filled out. The autopsy permission agreement signed by the individual is for an initial period of 5 years; the USUR ordinarily seeks renewal of the agreement upon its expiration. While this agreement provides

Jan 1, 1981

By R. K. Mehta, C. W. Schultz, J. B. Bates

Introduction The Mineral Resources Institute, The University of Alabama, has for the past three years been engaged in a program to develop a beneficiation system for eastern (Devonian) oil shales. One objective of that program was to evaluate advanced technologies for effecting a kerogen-mineral matter separation. Column flotation was among the advanced technologies selected for evaluation. Early in the program it was shown that column flotation was superior to conventional (mechanical) flotation and to the other advanced technologies being evaluated. The investigation then proceeded toward the further objective of defining the optimum operating conditions for column flotation. One observation made in the course of optimization testing was that introducing the feed into the froth (above the pulp-froth interface) resulted in an improved combination of concentrate grade and kerogen recovery. This observation was reported in a previous paper (Schultz and Bates, 1989). Because the practice of maintaining the pulp froth interface below the feed point is contrary to "conventional" practice, it was decided to subject the observation to a systematic series of tests. This paper describes a recent series of tests and the results that were obtained. Experimental equipment and procedure The arrangement of the column cell and auxiliary equipment for continuous flow testing is shown schematically in Fig. 1. The feed sump [O] is filled with a sufficient volume of prepared sample to permit a large number of tests to be performed (typically 12). Past experience has shown this is necessary to control sample variability and variability in the size distribution resulting from ultra fine grinding. The feed slurry is maintained at about 20% solids and is constantly recirculated and stirred. The sample is metered from the circulating pipe by a peristaltic pump [O]. The feed slurry is diluted with reagentized water [O] by a second peristaltic pump [O]. Wash water [O], also reagentized, is supplied through a third peristaltic pump [O]. While this feed system may seem unduly complex, it does permit users to independently vary either the wash water rate or the net solids content of the cell. In the tests reported here, the feed rate and net percent solids were constant at 12.5 gms/min. and 3.3%, respectively. Diluted feed enters the column through 6.35 mm-diam (0.25 in.-diam) copper tubing and is discharged upwardly at the center of the column. Tailings are discharged through flexible tubing that can be adjusted so as to control the position of the pulp-froth interface. The column is 76.2 mm-internal-diam (3 in.-internal-diam) and 1090 mm (43 in.) high. It is made from lucite tubing and is fitted with a 51-mm-diam (2-in.-diam) fritted glass air sparger having an average pore diameter of 50 µm. In performing a series of tests, the concentrate and tailing are allowed to discharge continuously. The system is allowed to equilibrate for 30 minutes after the pulp and froth reach operating levels. Concentrate and tailing samples are taken simultaneously for timed intervals (five to 15 minutes, depending on the volume of sample desired). After sampling, a change in operating conditions is made and the system is again allowed to equilibrate. The tests to determine the effect of the pulp-froth interface level were part of a larger series of tests in which the objective was to optimize the conditions for a rougher flotation stage in a two stage circuit. The sample used in this series of tests was an Alabama shale ground to d90 = 23.1 µm and d50 = 7.9 µm. The operating conditions remaining constant in this series of tests were as follows: Column height - 1600 mm (63 in.) Air sparser - 50 µm (average pore diameter) Spray water - 130 cc/min. Feed rate - 12.5 gm/min (0.4 oz per min) (dry solids) Percent solids - 3.3% Frother (Dowfroth 250) - 45 ppm The variable test conditions are tabulated in Table 1. Positions of the pulp level (pulp froth interface) and feed entry are presented as a percentage of column height (as measured from the face of the air sparser). These test conditions are presented Fig. 2. At each of these test conditions, individual tests were performed at varying air

Jan 1, 1992

By G. P. Larson

We define a specialized mineral market as follows: Specialized markets occur where a low volume of a given mineral is used to convey a large benefit to a specific product. Sales of these value added minerals to small markets generally require a different marketing approach than bulk or commodity markets for traditional minerals. Specialized markets are differentiated by several characteristics as listed in Table 1. [Table 1-Characteristics of Specialized Markets 1. Long-term sample evaluation 2. Small volume markets 3. Highly controlled properties 4. Customized specifications 5. Technical selling effort] Many of the specialized markets require long term evaluations before sales can be made. This can vary greatly depending on the type of industry and ultimate end-use of the final product. In some cases, complete evaluations may take from several weeks to years. Length of sample evaluation has no relationship to size of market. That is, the total market may be quite small or fairly large (e.g. from a few hundred pounds to a few million pounds). Uniformity of value added minerals from lot to lot is an extremely important characteristic for these markets. Variations that are of no consequence to one industry (or application) cannot be tolerated in another. This has led to customized specifications and test procedures for specific markets and applications. Because of the greater uniformity and customized specifications, the need for greater technical selling effort is required. This is particularly true in the initial developmental phases and contacts as well as in maintaining the account after the initial order. A small entrepreneurial organization can be used to advantage to achieve goals of sales and profit to these types of markets. The large commodity or bulk producer's sales and marketing organizations are not ordinarily equipped to give the necessary service or technical assistance to the specialized markets. The organization usually required for this type of marketing is given in Table 2. [Table 2- Specialized Marketing Organization Requirements 1. Flexible approach to technical and marketing problems 2. Versatile background in a broad range of industries 3. Laboratory backup when necessary 4. Marketing expertise 5. Technically competent management] In developing specialized markets, a flexible approach to both technical and marketing problems is necessary in understanding the needs and requirements of the industry and the applications involved. This flexibility is also helpful in adapting existing products to new market areas. Having knowledge of a range of industries as broad as possible can be very helpful. It often happens that a product developed in one area or industry can lead to its application in a different area with excellent results. Laboratory assistance for a wide range of tests and test procedures is an important asset in any development program for specialized markets. Laboratories can also assist in competitive evaluations. Marketing expertise and a management with some degree of technical understanding are necessary in to assign proper cost/ benefit ratios to a proposed new product or application of existing products. This is important since the total market volume can be small and the time required for initial contact to first sale can often be lengthy. Also, management's technical and marketing expertise is required in evaluating all possible options that can be considered viable when an application is presented. Once an organization of this type is in place, it is then possible to determine the needs of the marketplace, including when various minerals might have to be customized to fill those needs. There are generally two methods of marketing, as noted in Table 3. [Table 3- Method of Marketing 1. Product Driven -Develop a product with certain desirable properties and try to find a market to utilize those properties. 2. Market Driven -Know market well enough that when a need arises, it is possible to identify a product that will satisfy the need.] These approaches are not mutually exclusive and can be used at the same time. Marketing a value added mineral generally requires the "market driven" approach. Hence, the importance of a knowledgeable management as outlined above. To determine where needs and opportunities exist, several approaches can be used. Table4 lists some of the methods employed.

Jan 1, 1993

By N. Sheikh

In galena flotation, lead ethyl xanthate, Pb(EtX)2, is precipitated and is physically adsorbed in the form of multilayers on the chemically adsorbed first layer of Pb(EtX). 1,2) The solid particles are coated with layers of Pb(EtX)2 in heterogeneous manner over surface areas possessing both anodic and cathodic character (3,4) It is anticipated that the mode of formation of Pb(EtX)2 and the number of ions present, together with stable or unstable by-product species, have great influence on the surface characteristics of galena. The information obtained from a study of these parameters may have important practical applications in flotation. Lead ethyl xanthate is sparingly soluble in water. The metal and xanthate ions released to the aqueous solution during its dissolution undergo various types of reactions depending upon the pH values of the system. Extensive work (5,6) has been carried out to elucidate the mechanism involved in acid decomposition of xanthate ions. Different decomposition paths (7,8) have been proposed for the hydrolytic decomposition of xanthate in neutral and alkaline solutions, The concentrations of lead ions vary due to a number of chemical reactions, regulated by the pH and the temperature of the bulk phase. The hydrophobic character of Pb(EtX)2 may change with the nature of chemical reactions. Some of the numerous by-product species in solution may adsorb on the solid and change the properties of surface film either favourably or adversely. The flotation performance of galena, therefore, may depend to a considerable extent on the physical and chemical stabil¬ity of Pb(EtX)2 layers. Since little work has been reported in literature dealing with the stability of Pb(EtX)2 and its correlation with the flotation of galena, attempts have been made In this study to evaluate the behaviour of freshly formed Pb(EtX)2 in aqueous systems at various pH values.

Jan 1, 1971

By M. Ramachandran, R. G. Reddy

"A direct route for the reduction of magnesium oxide to magnesium using thermal plasma processing, with methane as the reducing gas, has been systematically evaluated. A feasibility study of the reduction reaction has been performed on the MgO-CH4 system at different methane concentrations and temperatures. The effects of changes in the key process parameters, such as the oxide-to-methane ratio and plasma power, on the yield of magnesium have been systematically evaluated. A maximum yield of about 61 mol% Mg was obtained using power of 20.0 kW and a molar ratio of MgO:CH4=1:1. Phase, morphological, elemental and binding energy analyses were performed on the product powders to understand the reduction reaction. Surface oxidation was observed in smaller magnesium particles in all the experiments, while the in-situ sintering of magnesium rods was observed at a molar ratio of MgO:CH4=1:1.5. Magnesium carbide (Mg2C3) was observed in the product, and its mechanism of formation is proposed using phase equilibrium data and a predominance diagram.IntroductionMagnesium has an excellent combination of low density and high strength. These properties have increased worldwide demand for and use of magnesium. Two major processes are used in the commercial synthesis of magnesium metal. The first method is electrolytic reduction and the second is thermal reduction using Pidgeon process. These processes can be classified as traditional techniques, but there are numerous emerging technologies that aim at reducing the production cost of Mg, its environmental impact and energy consumption. Thermal plasma processing is one emerging technique that has great potential in all the abovementioned key areas of magnesium production.Traditional reduction techniques. Two major reduction techniques, namely electrolytic and thermal reduction, have been used in the commercial production of magnesium. Electrolytic reduction is typically carried out in electrolytic cells from magnesium chloride (obtained from seawater or well and lake brines) (Tripp, 2009). The following simplified reaction shows the reduction of magnesium chloride to liquid magnesium and chlorine gas:"

Jan 1, 2015

By Thomas Schwind, Benedikt Grau, Angelos Gakis

"In Crossrail’s Farringdon station, compensation grouting was successfully utilized to mitigate the surface settlements induced by the tunneling works. The main focus of this study was the platform tunnels, which were approximately 300m long each and were enlarged to platform size using sequential excavation method (SEM) from the existing TBM running tunnels, initially supported using sprayed concrete linings. The rear grouting injections were carried out at a typical distance of minimum 5m above the completed, fully strengthened shotcrete lining hence inducing some additional stresses. In-tunnel monitoring using surveying targets was performed systematically to ensure that no excessive distortion of the initial shotcrete lining occurred. 3D finite element analyses were utilized to back calculate the actual “effective” pressure that was imposed on the completed shotcrete rings against the additional measured in-tunnel displacements due to the compensation grouting episodes. Additionally an assessment of the theoretical, “maximum” allowable grouting pressures that would lead to overstressing of the lining is presented. INTRODUCTION As part of the back analysis performed for the SEM tunnels in Farringdon station, an investigation of the effect of compensation grouting on the two constructed platform tunnels was undertaken. The platform tunnels were 10.6m high by 11.4m wide and approximately 300m long each. They were enlarged to their final shape from the existing TBM tunnels using SCL methods. This study utilized 2D and 3D finite element (FE) analyses in order to back calculate the observed tunnel crown displacement, assessing the equivalent grouting pressures during the rear injection episodes of the concurrent grouting operations. The results provide a useful reference for the order of magnitude of the grouting pressure load cases that can be applied in the numerical simulation of similar cases. Part of the results presented in this paper, were derived by the second author as part of his Master of Science thesis in Technical University of Graz."

Jan 1, 2016

By K. A. Natarajan, K. Karthiga Devi

"A biodegradable flocculant was produced during growth of Bacillus megaterium. The major component of the bioflocculant was found to be a polysaccharide composed of some proteins. Fourier transform infrared (FTIR) spectra analysis revealed the presence of carboxyl and hydroxyl groups in the bioflocculant, and thermal characterization by differential scanning calorimetry (DSC) showed the transition and crystalline melting point at 90-105°C. The effects of bioflocculant dosage and pH on the flocculation of mineral suspensions were evaluated. The bioflocculant exhibited good flocculating capability on mineral suspensions and achieved flocculating efficiencies of 90 percent for kaolinite and 85 percent for hematite suspensions at a dosage of only 5 mL/L. The maximum arsenite removal was found to be 90 percent at a bioflocculant dosage of 2 g/L, which is better than traditional chemical flocculants. This study demonstrates that microbial bioflocculants have potential for application in environmental cleanup, such as in the flocculation of mineral fines and in the remediation of solutions that contain toxic heavy metals.IntroductionBioflocculants are biopolymers produced by microorganisms during their growths. The predominant components of a bioflocculant are extracellular polymeric substances such as glycoprotein, protein, cellulose and nucleic acid (Salehizadeh and Shojaosadati, 2001). Bioflocculants have gained much wider attention due to their biodegradability and safety. In recent years, the study of bioflocculants has attracted wide attention. Bioflocculants can be nontoxic, harmless and eco-friendly. Normally, they exhibit different functional properties, serving as stabilizer, dispersant or thickener. Among these properties, flocculation is of major interest. For example, Phormidium sp. strain J-1, isolated from a drainage channel, exhibited excellent flocculation activity for bentonite particles in liquid suspension (Fattom and Shilo, 1984). A protein bioflocculant produced by Rhodococcus erythropolis could efficiently flocculate suspended solids while losing their flocculating activity by enzymatic digestion (Kurane et al., 1994, Takeda et al., 1991). Klebsiella sp. strain S11 screened from activated sludge was unable to flocculate without the addition of calcium chloride (Dermlim et al., 1999)."

Jan 1, 2015

By Fatima-Zahra Houmymid, Elena Chiriotti, Mahdi, Samy

"The Greater Paris projects involves the creation of a 200km extension of the actual public transport network on the close outskirt of Paris. In 2011, SYSTRA has been awarded the Engineering studies of a 20km underground section in the densely urbanized southeast quarter of the Parisian suburb. In order to both guide the TBM technology choice and identify and mitigate the risk due to soil settlement during tunnel boring, several analysis have been carried out and integrated into a Geographical Information System to share and manage this risk during the successive design phases of the project. The geological context‘s complexity and the Owner‘ requirements have led to the necessity of using an innovative analysis approach, leveraging both the numerical modelling analysis’s accuracy and the analytical method‘s ease of programming and implementation in a GIS, by the mean of the calibration of the analytical approach’ parameters on the FEM results.INTRODUCTIONThe Greater Paris Transport Project foresees the extension of the actual public transport network to the close outskirt of Paris through the extension lines 11 and 14, the creation of a circular new line around Paris - line 15 completed by lines 16 and 18 linking developing territories - and line 17 to Roissy-Charlesde- Gaulle airport.The SOCIETE DU GRAND PARIS (SGP) is the public institution that promotes this project and awards SYSTRA as Engineering for the 15th Line southeast section.Line 15th – South-East sectionIn 2011, the Owner SGP (SOCIÉTÉ DU GRAND PARIS) awarded to SYSTRA the preliminary and detailed design of a 20km-long underground section of Line 15, between the stations of Villejuif-Louis-Aragon and Noisy-Champs. The double-track, single-tube tunnel has an inner diameter of 8.5m (about 9.8m excavated diameter, 75m2 excavated cross-section). The tunnel is inserted in a densely urbanized context."

Jan 1, 2016

By Lukas Walder, Roland Arnold, Raphael Zuber, Richard Gradnik

This paper presents the construction of the ventilation shaft for the single-tube, doublelane Kramer tunnel. The project is located in the Federal State of Bavaria, Germany. This road tunnel, which has been under construction since December 2019, is the key component of the western bypass of the town of Garmisch-Partenkirchen. Once completed, it will relieve the town’s central district from enormous volumes of traffic. The contractor is the joint venture named ARGE Kramertunnel consisting of BeMo Tunnelling and Subterra. The whole project will be completed and handed over to the project owner, Staatliches Bauamt Weilheim, in 2023. The opening of the entire bypass section including the tunnel to traffic is planned for 2024. The tunnel, with a total length of 3,609 m (2.24 miles) and an estimated cost of EUR 178 million incl. VAT (US-Dollar 201 million), is designed as a single-tube, double-lane structure with one traffic lane for each direction of travel. The safety issue that exists for singletube tunnels is solved by a parallel escape tunnel passable for vehicles. For tunnel ventilation, an intermediate shaft has been designed with a depth of approximately 90 m (295 feet) and an inner diameter of 4.5 m (14.8 feet). All tunnels and ducts are excavated according to the principles of the New Austrian Tunnelling Method (NATM). The geological and hydrogeological conditions, the change in the profile (from round to square geometry), and also the shaft being located at a high elevation presented a demanding technical and logistical challenge. The site arrangement with the pit bank was located in a cut, the slopes of which had to be stabilized with shotcrete and soil nails. This excavation was carried out from the fall of 2020 to the beginning of 2021 because construction operations had to be completed outside the nesting period of protected birds.

Jun 13, 2023

By Kenneth P. Katen

Though the use of continuous mining machines consolidated the operations of cutting, drilling, blasting, and loading in one machine that would theoretically provide uninterrupted production, it was soon realized that the face haulage and roof support operations remained as major impediments to continuous production. The ability to provide not only continuous cutting and loading, but also continuous haulage and roof support (see Fig. I), was the basis for the introduction of mechanized longwall systems in this country from Europe in 1951. The early systems employed friction props for roof support, but were to become obsolete shortly after introduction of the first hydraulically powered, self-advancing supports in 1960. Since then, use of the longwall technique has grown to the point where it is currently providing more than 5% of the annual output of underground coal mines in the United States. The shortwall system had its origin in Australia, but was first introduced into this country in 1962 following the development of powered roof supports. It, too, provides for continuous cutting, loading, and roof support (see Fig. 2), but has not experienced the widespread popularity of longwall due to its inherent inability to easily provide for continuous haulage. While there are in the neighborhood of 100 longwall systems currently in operation in this country, there are only a handful of short- wall operations providing production of slightly more than 907 184.7 t/a (1,000,000 stpy) Both of these systems provide uninterrupted roof support for face operations by means of hydraulically powered steel roof supports. In each case, the coal removal consists of open-end cuts made from the narrow end of a large rectangular pillar (panel) which has been blocked out by continuous miners developing entries around its perimeter. However, this is the extent of the similarities between the two systems. It is hoped [ ] that the information presented in this chapter will enable the reader to better understand the two mechanical systems and the mining concepts involved in their employment.

Jan 1, 1981

By Douglas G. Hubert

To put coal in its proper perspective as a source of energy for electric power production, we must first compare it to other energy sources. A report of the Atomic Industrial Forum dated February 1958 records a number of authoritative though widely varying predictions. The most optimistic of these states that by 1968 there will be a total U.S. capacity of about 6000 megawatts of nuclear power, if growth continues at the present rate. Other sources indicate over 200,000 megawatts of total electrical capacity in utility plants by 1968. This is only 3% nuclear. Therefore, 97% of our total power production will be from non-nuclear energy. Fig. 1 shows the distribution of energy sources of electric utility product- ion in the U.S. as reported by the Federal Power Commission in 1954. Note that less than one quarter of the total is hydro and about one half of the total is coal. For 1957, the Edison Electric Institute reports that of the total steam- electric generation in the U. S., 68.6% was by coal, 7.9% was by oil and 23.5% was by gas. This is an increase for coal and a decrease for oil and gas in those three years. Thus we can conclude that two-thirds of the steam-electric power in the U. S. is now generated by coal. Fig. 2 shows the major markets for coal for 1947 vs 1957. The greatest in- crease has been in the electric utility market, where it has more than doubled in the ten years. One-third of the total coal produced goes to the utilities which now represent coal?s major market. Utilities used 160 million tons of coal in 1957, and the estimated quantity for 1967 is 275 million tons, and for 1977 is 525 million tons. It is predicted that 85% of steam-electric power will be generated by coal in 1975.

Jan 1, 1959

By M. C. Fuerstenau

The zero-points-of-charge of apatite, calcite, and fluorite are pH 6.4, 10.8, and 10.0, respectively. Scheelite is negatively charged above at least pH 3. Flotation responses of these minerals in the presence of potassium oleate and sodium silicate are presented and compared with electrokinetic data. Colloidal silica appears to be the species principally responsible for calcite depression, while silicate anion is the species responsible for fluorite depression. Additions as high as 1 x 10-3 mole/liter silicate have no effect on the flotation responses of apatite and scheelite. Selective flotation of nonmetallic minerals is difficult to achieve with fatty acids or soaps by themselves. As a result, specific reagents are added to aid these separations, and one of the reagents commonly employed for this purpose is sodium silicate. Flotation separations of various calcium¬ bearing minerals such as fluorite from calcite (1,2,3) and scheelite from calcite(24) and apatite(2,5) almost always involve the use of sodium silicate, for example. The mechanisms by which sodium silicate functions as a depressant are still not understood; this fact is probably due to a number of reasons. That is, the dissolution process of sodium silicate is complex, giving rise to a number of ionic and colloidal species. (6) In addition, the type and concentration of these species depend on the ratio of Na20 to Si02, the concentration of sodium silicate, and the pH of the system. (6) At the present time, it is not known which species, colloidal silica or silicate anion, is responsible for depression. If colloidal silica is the species that is adsorbing, then adsorption must occur by electrostatic attraction between the colloid and the mineral surface. Silicate anion, on the other hand, may adsorb either physically or chemically. The objective of this paper is to determine first the active species of sodium silicate and then the conditions under which this species will adsorb and function as a depressant.

Jan 1, 1968

By Peter Zhang

The immediate roof at the test site consisted of laminated clay shale that is likely to be susceptible to weathering. With seasonal dry-wet cycles in the mine ventilation air, the roof in the intake air would deteriorate and lose integrity over time. In many instances, the weathering-prone roof would gradually scale off piece by piece between bolts and over the entry corners. In some long-term entries, roof weathering could develop up to 2-3 ft deep with columns of bolted roof hanging under the upper intact roof. If wire mesh is used, the roof bags down and reduces the entry height, and the weight of the bagged roof tends to pull the bolts out and poses a safety hazard. An effective measure is always sought to support this type of weak roof that is sensitive to moisture degradation. One solution is to seal the roof shortly after entry development to prevent moisture penetration so that roof integrity can be maintained for the long term. In this application, a polymer-based sealing material was tested in the Rockspring mine for weathering protection and reinforcement. The material is a one-component polymer powder for spray application onto the roof and rib. The sealant material is mixed with water in the spraying nozzle to apply to the roof and rib as a paste. The product sets within 5-10 minutes and forms a membrane as thin as about 1/8-in. The membrane will quickly develop tensile and bonding strength within minutes and completely cure within a few days. Through sealing, bonding and lateral reinforcement, the integrity of the weak roof can be maintained. The material has been tried in a long-term evaluation over several months in a set of main entries in the Rockspring mine owned and operated by Rockspring Development, Inc., affiliate of Foundation Coal. Preliminary evaluation has shown that the material is an effective and economical solution to ensuring the long-term stability of weathering-prone roof.

Jan 1, 2009

By D. W. McGlashan

[Becaune ad ;the extennive nwrgace anea od u.X chopRex2 in emmiom, emuhion dRokdtiun oddeh poaaibWa don ;the necove/ry ul; BinQey divided minmcd pa~LL&a, Swr~aC;tavu% mu2 akab&ze ;the emu-bion and poan an an ad~i&y don ;the dahed mina&. OZhm hpohtant acto oh me ;the naAwre od ;the em&iun inghedied, ;the nQRaLLve volwna 06 the colz;tinuoun and dincon;tinuoun phaa, ;the change aign bonne by the dnuplm ad the o.X phae, ;the z&a pu;tenticd od ;the dahed minmd, the ~olubRe am curdent 06 khe wcr;ten-pulped and gnuund onu. AX on nem ;the AoQRec;trtic point on ZPC, ;the doubRe Raym about ;the minW-&den oil dnopR& i~ neduced in ;tkichnaa and the dn_opR& wU codace and continue to aggnegate into agglommaka . Invetmion dnom O/U ko W/O can UCCWL depending upon a nwnbm a 6 in;tmQeaked daotuh~. EReothokin&Lc inn&wnenta;t/iun to meanwre ;the ZPC, QReckProphonQkic mobilXty z&a pokeW and hakerning cuttrtent me invcduabte in ;the appficatiun ud emu-bion dRo;tat;iun ;to m&~c and nonm&~c minu.&. ] INTRODUCTION Recent interest in emulsion flotation can be ascribed to the necessity of selective separation and concentration of extremely fine minerals. That which follows is an elucidation of the many factors-in emulsion technology in application. In applied surface chemistry for solid-particulate separation and concentration, emulsion flotation is one of the intermediate processes between froth flotation, in which the continuous phase is water (99.5 to 99.999% by volume), and particulate extraction, in which the continuous phase is almost entirely an organic-fluid. Another intermediate process is particulate exchange (or partition) in which immiscible or very limitedly miscible phases are in controll¬ed but appreciable volumes. Surfactants are common to all of these

Jan 1, 1968

By J. M. Ekenes, C. A. Caro

"New approaches that increase hydrometallurgical recovery from refractory minerals in below mill cutoff -grade sulfide copper ore can add significant value, improve profitability and increase ore reserves. This paper describes the development and implementation of an engineered heap (EH) approach to leaching low-grade, chalcopyritebearing ore at Freeport-McMoRan’s mine in Morenci, AZ. The technical evolution of the engineered heap concept is described together with the challenges and benefits following implementation. Introduction The Morenci ore deposit, located in eastern Arizona, has been mined for copper for 140 years. Mineralization in the Morenci district consists of sulfides and oxides. To process these ores, Morenci operates concentrator and leaching circuits. Leaching accounts for two-thirds of the copper produced in Morenci. More than one-half of this leaching-derived copper production comes from run-of-mine (ROM) stockpiles. The largest category of material moved in Morenci is destined for ROM stockpiles and waste dumps. Materials that are not economic to mill or leach are designated “waste.” However, this designation can change based on improvements in extraction technology and commodity prices. Chalcopyrite does not leach well due to its refractory nature. For predominant chalcopyrite mineralized ore, copper recovery during a typical leaching cycle may be as low as 5%. For this reason, chalcopyritedominant ores below mill cutoff grade have historically been classified as waste. By incorporating advanced leaching technology to improve metal recovery from previously leach-resistant ore types such as chalcopyrite, the Morenci Mine has been able to reclassify substantial amounts of this “waste” to “ore.” Process chemistry Chalcopyrite copper extraction is temperature, oxygen and ferric dependent. The temperature is achieved primarily by pyrite oxidation (Eq. (1)) and secondarily by oxidation of chalcopyrite itself (Eq. (2)). The regeneration of ferric in solution can be enhanced by providing the appropriate acid and oxygen levels according to Eq. (3). In the absence of microorganisms, this reaction is very slow."

Jan 1, 2015