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By W. J. Wolfe

Using a combined fire assay-atomic absorption method, gold determinations with a minimum detection limit of 5 ppb were made on 345 samples of Early Precambrian felsic plutonic rocks collected 2 2 from intrusions ranging in size from 4 km to 88 km and in composition from syenite to granodiorite. Gold levels in unmineralized stocks are consistently below the 5 ppb detection limit of the analytical method used. In contrast, analyses of samples from stocks containing gold mines within their borders and in nearby metavolcanic-metasedimentary host rocks show a strongly skewed trace element gold population with a substantial percentage of gold values exceeding 5 ppb and ranging as high as 320 ppb. The data indicate that (1) ore-stage hydrothermal metasomatism has produced widespread Au enrichment in Precambrian intrusions that host epigenetic vein-type gold deposits, and (2) geochemical Au analysis of samples collected from 30 to 50 randomly distributed sites can be used to estimate the gold exploration potential of stock-sized bodies and to outline internal patterns of Au variation which may be useful in directing exploration to particular contacts or portions of a stock. The results do not conclusively indicate unusually high Au content in primary unaltered igneous rocks.

Jan 1, 1975

By Michael L. Clar

Reclamation of surface mined lands has two basic objectives: (1) the elimination or abatement of undesirable environmental conditions, and (2) the restoration of mined land to its former or more productive uses. This paper addresses some of the planning concepts and site factors that influence the selection of suitable land uses for surface mined lands. Three planning concepts included in the discussion are: (1) mine land planning, (2) comprehensive land use planning, and (3) environmental site planning. The site factors that will influence the selection of an appropriate land use have been grouped into three categories: (1) legal, (2) geographic, and (3) natural. The natural factors are further subdivided into three groupings: geomorphology, soils, and climate.

Jan 1, 1981

By Rusdian F. Kayadoe, Michael B. Hutahaean, Sandi A. Firmanulhaq

Deep Mill Level Zone (DMLZ) mine is located in Papua, Indonesia operated by PT Freeport Indonesia (PTFI) with approximately 1,500m below surface. The ore body extend 1230m width from northwest to southeast by 300–400m length east-west. The mine currently produces at full production of 80,000 tons per day until 2041. Part of the extension of current active production block and mitigation sits 225 m above the production block 1 (PB1) and 2 (PB2A and PB2B) which located in 2590L, named production block#4 (PB4W and PB4E). As the mining continues at PB1N, PB2A, and PB2B, the cave growth vertically and laterally to south perimeter of the footprint affecting the integrity of PB4W/E footprint. Currently, PB1N and PB2N are in active production whilst PB4E development continues from South against PB1N cave, as can be seen in Figure 1. A slight overlapping between PB1E and PB4E intentionally to promote cave connections between the two caves. Despite low stress model in the PB4E near existing caves, the PB4E cave propagation expected by the large strain and will be propagating in tension failure to the PB1E cave. Hence, designing the drill drives edge and first drawbell location as close as possible to the PB1E cave is necessary to guarantee cave connection. Since PB4 footprint development commenced in December 2021, there have been a substantial cave propagation observed from existing downhole monitoring. As the PB4 footprint designed perpendicular to existing DMLZ PB1 cave, adjustment of the development edge design is necessary to reflect the progression the loosening zone propagation to the South. Final life of mine design for extraction, undercut and ventilation level were decided in October 2023 considering latest update cave (fractured/ yielded) boundary. Numerical model of potential rock mass damage also aids the decision to locate the drift edge design. Predicting the cave progression impact is critical to both production at the existing active blocks and development of the new area. The use of cave shape data from year 2021–2023 gives an insight of the cave shape estimation at the end of 2024 when the undercut drift will be developed. This paper presents actual cave growth of active production blocks and its relative vertical progression impact as the result of drawing in order to plan for acceptable rate and provide mitigation once the minimum acceptable rock mass condition and minimum pillar threshold achieved.

Feb 1, 2025

By Thomas L. Craig

The Wellman-Lord SO2 Recovery Process is operating commercially at one location in this country and at two locations in Japan. Tail gases from a sulfuric acid plant, from an oil fired steam generator and from a sulfur recovery unit are the respective applications. 1) At the Olin Corporation's Paulsboro, New Jersey, stack gas from a spent acid regenerative sulfuric acid plant-45000 SCFM - is treated in the W-L unit to reduce sulfur dioxide concentrations from 4500 ppm to less than 500 ppm. 2) At the Japan Synthetic Rubber plant in Chiba Prefecture, stack gas from an oil-fired steam generating plant - 120,000 SUM - is reduced from 2100 ppm S02 to less than 200 ppm. 3) At the Toa Nenryo Kogyo refinery in Kanagawa Prefecture, tail gas from a sulfur recovery (Claus) unit - 41,000 SUM - with, an inlet SO2 concentration of up to 13000 ppm is reduced to less than 200 ppm. The three installations are operating successfully and are capable of reducing S02 emissions to levels significantly below the concentration guaranteed. With the one exception of a higher than guaranteed chemical make up requirement at Paulsboro, utility and chemical requirements will be less than design.

Jan 1, 1971

By D. Kidd, T. Tucker, L. M. Indergard, S. Davis

"INTRODUCTION The Lisbon Valley Mining Co LLC (LVMC) operates the Lisbon Valley Mine located in SE Utah. The mine includes 3 open pits, 150-acre heap leach pad, and solvent extraction electro winning plant (SX/EW). Approximately 25MM tons of ore has been mined since 2005. The ore is stacked on the leach pad to an average height of 55 feet and occurs as a blend of crushed material (minus 2”) and run-of-mine (minus 24”). The host rocks include Dakota and Burro Canyon Sandstones. Copper occurs finely disseminated in this material at an average grade of 0.46%. Minerology is 63% sulfide, 27% oxide, and 10% insoluble. Total inventory mined through 2016 is approximately 230 MM lbs. METALLURGICAL CONSTRAINTS Copper production is constrained by areas of low permeability ore and geochemical issues, including acid consumption. The intrinsic permeability the ore ranges from 10-2 to 1 Darcy. This permeability does not permit sufficient aeration by natural convection, therefore oxygen transport is largely limited to the mass of oxygen that can remain dissolved in leachates. As a result, greater than 50MM lbs of copper remains in inventory and occurs secondary sulfides, including chalcocite, covellite, and bornite. BIOLEACHING FUNDAMENTALS Due to the higher activation energies of secondary sulfides, notably covellite and bornite, a strong oxidant is needed to remove an electron from the metal-sulfide lattice and release the copper cation into the sulfate solution. This process occurs most efficiently using the strong oxidant ferric iron (Fe3+). Ferrous iron can be oxidized to ferric iron by oxygen via chemical oxidation or biologically; which was measured to occur 500,000 times faster when using microbes versus oxygen.1 The microbes that oxidize ferrous iron (Fe2+) to ferric iron are naturally abundant and will develop biomass when appropriate conditions and nutrients present themselves. Most nutrients are provided by the ore/irrigation (P,N,Mg, Ca & Fe2+) while the others (O2 and CO2) are provided by air. In most cases, oxygen transfer is the rate limiting step during copper sulfide leaching. 1 Lacey & Lawson 1970 FORCED AERATION SYSTEM DESIGN This paper describes LVMC’s efforts to bioleach the remaining inventory by forced aeration. Two basic aeration systems were installed and tested including vertical wells and horizontal piping. Vertical System The vertical system was installed using the mine’s blast hole drilling rig. Wells were constructed using 4” high-density polyethylene (HDPE) connected to 4” polyvinyl chloride (PVC) screens similar to conventional bioventing wells. Ore samples were collected at each well location to characterize the spatial distribution of the remaining inventory along with fines (minus 200 sieve), pH (acid consumption), and iron. A block model of each analyte was developed and mapped in 3D."

Jan 1, 2017

By G. Castillo, J. Mortensen

"INTRODUCTION Molybdenite is typically produced as a co-product in the copper ore flotation process. At the Rio Tinto Kennecott Concentrator (RTKC) in Copperton, Utah, the molybdenite is initially floated with copper sulfides and other value minerals producing a Cu-Mo concentrate. This is then separated in the molybdenum plant by depressing copper sulfides and pyrite and floating only molybdenite. The molybdenum plant in RTKC comprises the rougher circuit and several stages of cleaning to achieve a certain grade requirement of Mo. The rougher circuit is where the first stage of separation occurs. Cu sulfides and pyrite are depressed in the rougher circuit by adding NaSH under nitrogen and in the cleaner circuit by adding NaSH under air. The depression of Cu sulfides (and pyrite) by NaSH is related to the redox activity1. It is important therefore to maintain a reducing potential in the circuit in order to achieve effective Cu-Mo separation. If the rougher circuit is under air, like RTKC’s molybdenum cleaner circuits, it is harder to maintain the target redox potential because NaSH oxidizes rapidly in the presence of air thereby increasing NaSH consumption significantly. CU-MO SEPARATION AND ROLE OF PULP POTENTIAL The collector adsorption on copper sulfide minerals is through an electrochemical mechanism2. The collector changes the surface chemistry of the Cu mineral allowing it to float. Molybdenite, or MoS2, on the other hand, floats readily in the presence of a frother as it is naturally hydrophobic. Usually, hydrocarbon oil extenders such as diesel or burner oil are added to enhance Mo recovery. In order to effectively separate Cu from Mo, the collector complex on the copper surface needs to be destabilized or desorbed using a reducing agent such as NaSH or Nokes. As long as the circuit is under highly reducing conditions, the collector remains detached from the sulfide mineral surface allowing Cu sulfide depression. Therefore, it is important to measure the oxidation-reduction potential (ORP) continually and maintain the reducing condition with addition of NaSH. For a given collector each mineral will have a characteristic potential or electrochemical window for flotation to occur3, therefore, it is necessary to reduce the potential outside of this window in order to effectively separate Cu and Fe sulfides from MoS2 minerals2. In RTKC plant, the target reducing condition is between -450 to -550 mV to effectively destabilize the collector and depress Cu and Fe minerals. Typically the dosage of NaSH used in the molybdenum plant could range from 6 to 30 gallons/min (gpm). The high dosage of NaSH could pose risk of H2S exposure. RTKC wanted to evaluate alternatives to NaSH in order to reduce its consumption and minimize risk of H2S gas generation."

Jan 1, 2016

By M. E. Ryan, C. L. Hoebbel, E. J. Haas

DISCLAIMER The findings and conclusions in this paper are those of the authors and do not necessarily represent the official position of the National Institute for Occupational Safety and Health, Centers for Disease Control and Prevention. Mention of any company or product does not imply endorsement does not constitute endorsement by NIOSH. ABSTRACT It has been suggested that effective disaster prevention and response in mining requires that elements of emergency management be incorporated into everyday operations, planning, and decision-making, much in the same way that the industry approaches routine risk management. Since Congress passed the MINER Act in 2006, the newly mandated structural and technological advancements in mine emergency response preparation have been made and are apparent, but measureable improvement in mineworker emergency preparedness is difficult to ascertain. To begin exploring whether measurable indicators of mine workforce preparedness can or should be addressed within a risk management framework, NIOSH researchers examined the datasets from two separate research efforts—one which measured the organizational safety climate and self-reported safety performance, and the other which measured individuals’ perceptions of their own self-escape competency. INTRODUCTION In response to the mine disasters in 2006, the United States Congress passed the Mine Improvement and New Emergency Response Act (MINER Act), which strengthened existing regulations [30 CFR, Part 48] and introduced new provisions to improve mine emergency response. Since that time, new technological and structural features, such as mandated redundant communication systems, personnel tracking systems, and lifelines in primary and secondary escapeways, have been installed in underground coal mines across the country. Additionally, mine operators are required to develop emergency response plans, provide adequate supplies of breathable air throughout the mine, provide more frequent and realistic expectations training, and are required to immediately notify the Mine Safety and Health Administration (MSHA) in the event of an accident. While the structural and technological advancements in emergency preparedness are readily observable and easily measured, levels of individual mineworker preparedness remain difficult to assess. There is a general consensus that deficiencies in competence in the non-routine tasks required for individual miners to effectively respond to a mine emergency likely remain (e.g., National Research Council [NRC], 2013), and NIOSH research continues address this concern (Haas, et al., 2015; Hoebbel et al., 2018). To maximize an operation’s capacity to effectively prevent disasters and/or respond to emergency events, it has been suggested that individual and emergency management efforts be incorporated into everyday operations, planning, and decision-making in much the same way organizations manage routine risks to health and safety (Haddow et al., 2017). To the authors’ knowledge, the mindful intersection of routine risk management policies and practices and emergency prevention and response readiness have yet to be empirically studied within the mining industry.

Jan 1, 2019

By Andrew N. Peterson

US mine operators frequently fail to investigate more cost effective and productive bulk material handling systems because surface mines seem to lend themselves to truck ore haulage. In this country, as a result, use of conveyors to move heavy loads from mine to process facilities has been minimized, if not actually neglected. In contrast, there are more than 50 conveyorized surface mines in successful operation around the world. These mine operators have learned that properly applied conveyorized systems can offer major savings in capital and operating costs, which contribute to improved profits when combined with other proven mining technologies. Growing acceptance and application of conveyorized bulk material handling in surface mines also points up how unique each mine is and how careful planning contributes to maximum mine effectiveness. Because of these differences, mining executives and technical and operating staffs need to develop an understanding of three factors in applying conveyorized bulk material handling in surface mines: • Why each mine will benefit from the type of automation permitted by conveyorized operation, •What kind of equipment is available, and • What applications most effectively demonstrate the first two factors in action - hauling either ore or waste. The conveyorized systems considered in this presentation have production rates from 0.5-2.7 kt/h (500-3,000 stph). Worldwide, these systems have been operating since the early 1960s. Advantages of Conveyors Why do you want conveyorized bulk material handling? First, it almost always provides lower operating and maintenance costs. Second, it frequently requires lower initial capital costs and almost always requires lower capital costs over the life of the surface mine. Third, it provides comparable operating availability, and finally, it frequently gives comparable operating flexibility - depending on the mine plan. Cost avoidance can be accomplished with modern production methods. These, in turn, permit increased productivity and reduced operating costs such as those for energy, maintenance, and manpower. It has been demonstrated in European surface mines and elsewhere, that conveyor systems frequently require lower initial costs than does truck haulage. Almost always such operations require lower capital costs over the mine life. Those costs include the continual addition of haulage trucks to both accommodate the increasingly difficult haulage routes and fulfill replacement requirements when trucks wear out. Conveyor systems handling ore in numerous large crushing and port facilities, which have operated since the early 1950s, have clearly demonstrated a useful conveyor life of more than 25 years. In contrast, off-highway trucks have life spans of six to eight years. The following examples illustrate comparative capital costs to purchase conveyor systems and comparable truck haulage units. Example 1 The ore haulage route from point A to point B is level and 610m (2,000 ft) long. The material weighs 1.8 t/m3 (110 lbs per cu ft) and must be transported at a rate of 1.8 kt/h (2,000 stph). The installed capital costs to provide a properly designed conveyor that will transport the described material from point A to B is about $450,000. The capital cost to purchase three 77-t (85-st) off-highway trucks and one spare truck - which would provide equivalent capacity - would be about $1.2 million. The truck cost estimate is based on a 6 min. or 771 kt/h (850 stph) truck cycle time. Truck efficiency is estimated at 0.8. Each 77-t (85-st) truck would have an actual haulage rate of 617 kt/h (680 stph). Therefore, three trucks would be necessary to transport the designated tonnage of 1.8 kt/h (2,000 stph). A movable crushing plant would be located at point A for the conveyors and a permanent crushing plant at point B for the truck haulage system. Capital costs for these primary crushing plants were not included in the calculations for either system because the capital costs are frequently comparable. Example 2 The transport route from point A to point B is 610 m (2,000 ft) horizontally and 122 m (400 ft) vertically - on a 20% grade (Fig. 1). The material weighs 1.8 t/m3 (110 lbs per cu ft) and must be moved at a rate of 1.8 kt/h (2,000 stph).

Jan 6, 1983

By Kiseong Hyeong

Geochemical characteristics, shapes and surface textures of manganese nodules from two blocks located in the southern (9°N~11°N) (KR5) and northern parts (16°N~17°N) (KR2) of the Clarion-Clipperton fracture zone between 130°W and 133°W were investigated in terms of sediment distribution and metal contents for better understanding of controlling factors of seabed manganese nodule genesis. Two blocks, ~700 km away, are dominated by the nodules with different origins, each of which has distinctive chemical compositions and shapes. KR2 block is dominated by irregular and poly-type nodules of hydrogenetic origin with low Mn/Fe, Cu, and Ni contents and high Fe and Co contents, while KR5 block is dominated by discoidal and ellipsoidal nodules of diagenetic origin with high Mn/Fe, Cu, and Ni contents and low Fe and Co contents. In addition, nodules from two blocks show different surface textures: smooth and transitional in KR2 and dimorphic and rough in KR5. Sediment distribution is also distinctly different in two blocks. The study area consists of three sediment types: Unit 1 ? dark brown mud with an intermediate metal content, Unit 2 - yellowish brown mud with poor metal content, and Unit 3 ? black metalliferous mud. KR2 is characterized by the thick occurrences (<250 cm) of Unit 1 at the surface while Unit 1 distributes thin (<12 cm) in KR5. Higher primary production rate in the southern block proximal to equator and resulting high supply rate of metals to the sediment might be the cause for the dominance of diagenetic nodules in the southern block, but potential controls of different sediment distribution on the nodule genesis can not be excluded. If it is the case, Unit 1 with an intermediate metal content likely has been acting as a sink for metals by precipitating micro-nodules within the sediment column. Within the KR5, the hydrogenetic nodules occur preferentially on the seabed where Unit 3 distributes at shallower depth (<50 cm). Unit 3 has 2 or 3 times higher metal contents than other types of sediments. It might play the same role as Unit 1 in the northern block that prevents the supply of metals to the sediment surface. It also supports the potential effect of sediment distribution on nodule genesis. Keywords: Nodules genesis, Sediment distribution, Supply rate of metals

Jan 1, 2008

By S. E. Phillipson

The MSHA Technical Support, Roof Control Division's primary evaluation tool for coal pillar designs is the Analysis of Coal Pillar Stability (ACPS) program. However, ACPS is limited to evaluating interactions between two seams at a time, which is not ideal for complex multiple- seam scenarios. Further, it does not recognize the influence of strata properties in the interburden and cannot be used to evaluate individual pillars. We apply the Rocscience RS2 two-dimensional finite element program as a complement to ACPS. Complex multiple-seam scenarios can be characterized by a modified pillar stability factor where the ACPS single-seam development pillar stability factor is multiplied by the ratio of average single-seam development stress to the stress in the critical pillar at the multiple-seam stage, using the values of o as predicted by RS2. Several series of test examples involving two-seam analyses with isolated remnants, gob-solid boundaries, and high extraction areas of varying width were used to calibrate simulated gob properties such that the ACPS+RS2 method generated results that matched a traditional ACPS Stability Factor to within 13%. As high-extraction areas become wider and deeper, progressively stiffer gob is simulated by increasing the Young's Modulus property in RS2. After achieving calibration with the test series, three examples of complex multiple-seam scenarios were evaluated, using reasonable properties for strata types and thicknesses based on core log data. Resulting modified stability factors are reasonably close to those predicted by ACPS, although affected by units of sandstone in the interburden and gob shadows superimposed on remnant structures. This tends to show that even in complex multiple-seam conditions, applying ACPS to the worst-case interaction should usually result in an acceptable pillar design.

Jun 25, 2024

By Clifford W. Schultz

Four native copper ores, two conglomerate and two amygdaloidal, were tested to determine their amenability to sorting. One of them, a conglomerate ore from the Kingston mine, showed a high sortability as evidenced by an 80 pct recovery of copper in the minus 6-, plus 1/2-inch size range. Using the Kingston as a case example, the cost of mining and concentrating was estimated for a standard case and five optional cases involving varying degrees of sorting. This analysis indi¬cates that the production cost per pound of copper can be reduced on the order of 1 to 1-1/2 cents per pound along with a significant decrease in capital investment.

Jan 1, 1975

By Haydn H. Murray

Kaolin is an important industrial mineral because it exhibits desirable properties in many applications and because of its relatively low cost, Approximately 2,000,000 tons of kaolin from Georgia and South Carolina are processed and used commercially in the United States each year. Strictly defined, the name kaolin is applied to a group of hydrous aluminum silicates of which the mineral kaolinite is the most common. Kaolinite has such desirable properties as natural whiteness, chernical stability, non- abrasiveness, and very fine particle size. Kaolinite particles are characteristically thin, flat, and hexagonal in outline. The particles range in size from submicron to as much as 40 microns (equivalent spherical diameter). In addition to the general properties mentioned in the first paragraph, it is soft, has low viscosity at high solids content in many systems, is readily wet and dispersed in water and some organic systems, and can be produced with a controlled particle size distribution. Some of the important physical constants of kaolin are: specific gravity, 2.60; index of refraction, 1.56; hardness (Mohs scale), 2; fusion temperature, 1850°C; dry brightness, 78-92 percent.1

Jan 1, 1965

By Jr. Smith

Industrial minerals are just other rocks, except for the unique characteristics that make them useful in an industrial process. There is not a ready market for this type mineral and one or more uses must be developed to sustain production. The name "olivine" is a generic term used to indicate a group of orthosilicate minerals in a solid solution series, with forsterite (Mg2SiO4) and fayalite (Fe2SiO4) as endpoints. Since neither is found alone in nature, the mineral is always a combination of the two. The commercial deposits are high in magnesium and on the high end of the melting scale in the series. The name comes from the olive-green color.

Jan 1, 1992

By Salt Lake City Research Center, D. G. Foot, J. D. McKay

As part of its research program on column flotation, the Bureau of Mines, Salt Lake City Research Center has performed studies on (1) column flotation amenability and comparison with conventional flotation results for over 15 mineral systems, (2) alternative methods for improving column bubble generation systems, (3) mineral behavior within a column, and (4) the effect of changes in column operating parameters on flotation grade and recovery. Based on results from these studies, flotation columns (1) are capable of producing higher grade concentrates in fewer stages than conventional flotation cells at equivalent or increased recoveries, (2) require less retention time than conventional flotation cells, and (3) can be divided into cleaning and collection zones that have characteristic mineral-upgrading-per-foot values. Attempts to improve the column air sparger have led to the development of a superior external bubble generation system, which can (1) achieve increased air holdups. (2) produce bubbles from 400 Um to 4,000 Um average bubble diameter with minimal frother addition, (3) provide superior bubble size control, and (4) be maintained without column shutdown. Efforts are continuing to improve column flotation and transfer successful technology to industry.

Jan 1, 1988

By T. M. Gurr

The United States is the world&apos;s leading producer of phosphate rock. In 1975 according to the U. S. Bureau of Mines&apos; statistics, 44.3 million tonnes (48.8 million short tons) of phosphate rock were produced in the United States&apos;, which is 40.6 percent of the world&apos;s production2. The major phosphate producing areas of the United States are located in Figure 1 and the production for 1975 is listed as follows: Million Million Short Tonnes Tons Florida and North Carolina 36.9 40.7 Western States 5.3 5.8 Tennessee 2.1 2.3 All of these areas produce phosphate from deposits which are sedimentary marine in origin. Occurrences of phosphate exist throughout the United States in numerous locations that are of both igneous and sedimentary origins. However, these other phosphate occurrences vary greatly in extent and economic significance. Georgia, South Carolina, California, Texas, Arkansas, Kansas, Iowa, Alabama, Kentucky and Virginia are but a few of the states in which phosphate is thought or known to

Jan 1, 1977

By Enrique Fernández

INTRODUCTION The high speed rail link between Madrid and Asturias, in the North of Spain, has encountered in the Cantabrian range the biggest challenge of the whole line, a very deep tunnel, 25 km long through a foreseen difficult geology. This tunnel was divided in the following four lots for construction: ¦Lot 1: Pola de Gordón- Folledo. L=10,7 km ¦Lot 2: Folledo-Viadangos. L=4,166 m ¦Lot 3: Viadangos- Los Pontones. East tube. L=10,4 km ¦Lot 4: Viadangos- Los Pontones. West tube. L=10,4 km. Dragados S.A. leading a JV was contracted to design and built the Lot 2 on the captioned Pajares High Speed Railway Tunnel. The Lot 2, comprising the middle third section of the 25 km long tunnel has an additional challenge: the construction of one access from the surface to allow it own construction. Feasibility reports and preliminary design do not foresee the use of TBM due to the risk of squeezing ground at the deepest area of the tunnel. On the other hand, conventional NATM excavation of the twin tunnels and ancillary works, including access shafts were against the expected construction schedule by the client, due to the poor ground foresee at the early stage.

By VÍCTOR VÁZQUEZ-VÁZQUEZ, GUILLERMO TIBURCIO-MUNIVE, María M. Salazar-Campoy, JESÚS L. VALENZUELA-GARCÍA, OFELIA HERNÁNDEZ-NEGRETE

The present study examined a process for precious metal recovery (silver and gold) from difficult minerals which exhibit refractory characteristics to cyanidation and ammonium thiosulfate leaching (low extraction and high consumption of leaching agent). It is well known that mangano-argentiferous minerals are mainly composed of manganese, silver, gold, and iron. In this research, a mineral sample from Monte del Favor mine, located in Hostotipaquillo, Jalisco, Mexico, was physically and chemically characterized by X-ray powder diffraction (XRPD), scanning electron microscopy (SEM), energy-dispersive X-ray spectroscopy (EDS), fire assay, and atomic absorption spectroscopy (AAS). Two-stage leaching was carried out: first to eliminate manganese from the sample using sodium sulfite (Na2SO3) as a reductive reagent and sulfuric acid (H2SO4) and a second stage to dissolve silver and gold with sodium hypochlorite (NaClO) and hydrochloric acid (HCl); the results show extraction around 96% for Mn and 98.73% and 83%of Ag and Au, respectively. The usefulness of high acidity and Na2SO3 to obtain good results in Mn and Ag leaching is discussed along with the roles that HCl and NaClO play in the dissolution of Ag and Au values.

Apr 15, 2020

By Mahaboob Basha Shaik, Shakuntala Tudu, Indrajit Paul, Chandra Sekhar, Neha Gupta, Amit Kumar, Srinivas Dwarapudi, Arvind Kumar Patel

In this study, the effect of colemanite and calcined colemanite (CC) addition on the properties of iron ore pellets is studied. The pellets are prepared in a lab without and with 0.2% colemanite and varying amounts of CC from 0 to 1.0 wt%in steps of 0.2%. A reference pellet is prepared under typical plant conditions, and the properties of the reference pellet are considered as the base value for improving the cold compressive strength (CCS) while maintaining the other parameters at an acceptable limit without altering the pellet. The study also aims to provide a detailed characterization of colemanite and calcined colemanite to better understand their behaviour during firing. The lab-scale results show that the CCS of pellets made with 0.2% colemanite was reduced by 222 points (60%) as compared to the reference pellets due to crack formation during firing. The results show an improvement in the CCS of the indurated pellets with CC dosage up to 0.4%, and decreasing strength is noted thereafter. It is observed that the swelling index increases with the increase in calcined colemanite dosage. To simulate plant conditions, reference pellets and pellets with optimized CC dosage were fired in a plant furnace.

Oct 16, 2020

By M. Misra, I. A. H. Schneider

The flocculation of -38 pm calcite, hematite and kaolinite aqueous suspensions by the yeast Candida parapsilosis and products derived from rupture of the microorganism was studied. Flocculation of hematite suspensions by the microorganism and its derivatives readily takes place between pH 2.8 and 11.4. Maximum flocculation occurs at about pH 6.5. Flocculation of calcite suspensions by the microorganism and its derivatives readily takes place between pH 7.0 and 11.8. Maximum flocculation occurs at about pH 10.0. Flocculation of kaolinite suspensions was less than that for the other two minerals. However, a soluble fraction derived from the rupture of the organism functioned as a good flocculant for the mineral, albeit at an increased concentration, between pH 2.6 and 11.5. Maximum flocculation occurs between pH 6.8 and 7.2. An attempt is made to relate the behavior of the various systems to the surface characteristics of the minerals and the surface and chemical properties of microorganism and its derivatives. Experimentation used to evaluate the flocculation includes actual flocculation tests, microelectrophoresis measurements and scanning electron microscopy.

Jan 1, 1994

By L. Ou

This paper reports on research on leaching pyrite in a germ-free system. The effects of mineral particle size, pulp potential and various oxidation agents on pyrite leaching were studied. The results indicated that a smaller mineral particle size leads to a higher leaching rate of pyrite. The optimum results were obtained with a mineral particle size of -0.037 mm. The pulp potential reflected the leaching process and an increasing pulp potential improved pyrite leaching. The leaching rate rapidly increased with the addition of oxidation agents such as Fe3+ and H2O2. However, the precipitate generated by ferric ion coated the surface of the pyrite and prevented further leaching.

Jan 1, 2007