Search Documents

By Jaeheon Lee, Junmo Ahn, Jiajia Wu

To reduce the environmental footprint of hydrometallurgical processing of black mass from spent lithium-ion batteries (LIBs), a green leaching system based on glycine and sodium metabisulfite (Gly-SMS) was proposed. The leaching study demonstrated that the highest cobalt and lithium recoveries of 100 percent and 99.8 percent were achieved under optimal conditions. The leaching mechanism revealed that the dissolution of LiCoO2 in the Gly-SMS solution followed the shrinking core model. The apparent activation energies for cobalt and lithium were determined as 48.05 and 41.51 kJ/mol, respectively, indicating a surface chemical reaction controlling mechanism. The leachate was then processed by an acidification- precipitation technique with oxalic acid as the precipitant to remove cobalt. Glycine complexes with metal ions by zwitterionic ligand and recycles in the leaching-precipitation circuit, reducing the reagent cost. Compared to other studies, this leaching system has near-neutral operating conditions and is cost-effective, making it an economically viable alternative for treating cathode materials from spent LIBs.

Dec 1, 2024

By Dewa Putu Agus Sudarmaja, Widi Astuti, Januarti Jaya Eka Putri, Yayat Iman Supriyatna, Himawan Tri Bayu Murti Petrus, I Wayan Warmada, Siti Nurul Aisyiyah Jenie, Ferian Anggara, Anisa Maulidia, Vincent Sutresno Hadi Sujoto, Sutijan

With the rapid development of renewable energy technology, such as electricity reserves, to store energy for electric or electric vehicles, environmentally friendly energy sources are expected to continue to be fulfilled according to mobility and stationary energy needs. Lithium is the most attractive candidate for energy storage development. The presence of lithium in Indonesia in the Sidoarjo mud content is interesting to be used as an alternative source of raw materials. During the research, sulfuric acid acted as a solvent, and hydrogen peroxide acted as an oxidation agent. Lithium recovery will be studied using a leaching process at various temperatures (30, 45, and 60 °C) and sulfuric acid concentrations (0.5, 1, 2, 3, 4, and 6 mol/L). The increasing temperature, sulfuric acid concentration, and hydrogen peroxide concentration will directly affect the lithium recovery. An increase in sulfuric acid concentration will be directly proportional to the recovery of lithium in the solution. Optimum conditions were obtained at varying 6 mol/L sulfuric acid concentrations and 60 °C temperature variations. The shrinking core model (SCM) studied leaching kinetics, and ash layer diffusion acted as a rate-controlling process. The activation energy of the leaching process was obtained at 38.90 kJ/mol.

Jul 25, 2023

By Drummond Earley III

The hydrology of in situ leach mining operations is being studied by the U.S. Bureau of Mines using a variety of techniques, and the Bureau has made significant advances in tracking and modeling the flow of leach solutions in rock formations. However, the influence that leach solution composition has on the hydrology of these systems is poorly understood, and geochemically-controlled changes in the hydrologic characteristic of ore deposits during in situ leach mining of porphyry-hosted, copper oxide ores have not been previously studied. The Cyprus Casa Grande in situ leach mining operation mines fracture-hosted and disseminated copper oxides (predominantly chrysocolla) hosted by granodiorite porphyry using a pattern of underground wells. The injection solution consists of sulfuric acid raffinate from a solvent extraction-electrowinning plant. Chemical analysis of representative injection and recovery solutions reveals that gangue cation concentrations in these solutions are high, and geochemical modeling shows that the solutions are saturated or nearly saturated with respect to gypsum, jarosite, and various amorphous oxides. Precipitation of these solids is evoked by calcium-exchange reactions, rising pH, and perhaps locally rising redox potential, as the leach solution reacts with the ore minerals, host rocks, and the atmosphere. Mass and volume balance calculations show that approximately 25 cm of precipitates are formed per gram of copper recovered at Cyprus Casa Grande. The results of geochemical modeling agree with petrologic examination of pre-and post-leach core samples from the leached horizon. Moreover, hydrochemical modeling and hydrologic field data suggests that chemical precipitation causes significant decreases in porosity and permeability during in situ leach mining of porphyry-hosted, copper oxide ore.

Jan 1, 1992

By R. R. Klimpel

This paper describes a series of mechanical flotation cell results run with various chemical reagent schemes designed to economically improve the flotation performance associated with selected industrially available coals. The emphasis is on the identification of practical flotation methods of processing coals containing unusually coarse and/or fine feed sizes, of increasing the recovery of oxidized and/or difficult to float coal, and of increasing the selectivity of coal over pyrite in high sulfur coals.

Jan 1, 1991

By S. M. Moore

Each year more than 400 coal miners are injured (fatal and non-fatal) by rock falling from between or around roof supports. Many of these injuries could be prevented by the installation of roof screen (wire mesh). However, many coal mines may be reluctant to use roof screen because of the added cost. The goal of this study was to determine the potential saving in workers? compensation (WC) premiums that could be achieved due to a reduction in rock fall injuries after roof screening. The WC rate-setting methods utilized by IL and KY were investigated in this study. Using actual data obtained from MSHA, national and state WC bodies and individual insurance companies (e.g. average cost per injury, loss cost rate, number of injuries per year, number of injuries preventable each year with roof screening), hypothetical mines (representing two mine sizes: 67 and 150 employees) were constructed with realistic ranges for estimates of injuries and WC premium costs. Using each state?s actual WC rate-setting formulas, total savings in WC costs after a three-year period were determined. Across both states, savings in WC premiums ranged from 1.8% to 14.6% when injuries were reduced by 10% to 20%. Additionally, an economic analysis was performed for an existing mine consisting of 67 employees. For this mine, the annual cost of a roof screening program was estimated at $240,000. If the screening prevented 13 ?struck by? WC claims over a 3-year period at this mine, the reduction in WC premiums alone could pay for the entire screening program.

Jan 1, 2010

By Eugie Kabwe

The increase in depth of mines due to the depletion of surface deposits leads to unsafe tunnel construction and overall ore extraction. This is due to tunnel instability associated with the surrounding rock mass’s ability to exhibit creep that occurs over time. However, the explicit description of this mechanism using conventional constitutive models has proven to be difficult and inadequate. Hence, this paper presents an integrated viscoelastic viscoplastic visco-damage model with the Duvaut-Lions function. The presented model is calibrated by experimental data to determine the viscoelastic viscoplastic and visco-damage parameters. It is further implemented in a finite volume numerical code and verified on an experimental scale by numerical modeling of a cylindrical specimen. It describes the 3-phase creep mechanism perfectly as compared to the inbuilt conventional integer-order derivative model. Results from the conducted verification show that it can successfully be employed to describe the long-term behavior associated with the creep mechanism.

May 16, 2021

By F. Dehghani, M. Kebe, H. Askari-Sabzkoohi, T. Ghosh

The depletion of bauxite reserves around the world has necessitated the search for viable alternative sources of aluminum. The major non-bauxite source of aluminum is alunite [KAl3 (SO4)2(OH)6]. Lixiviation is the common industrial process to recover aluminum from alunite. Due to the presence of quartz (40%) as the main gangue mineral, this process causes the formation of Sodium aluminosilicate and Potassium aluminosilicate complexes that leads to loss of alunite, and chemical reagents. Thus, pre-concentration of alunite is required to remove the gangue minerals. Because of the low degrees of freedom (DOF) of alunite, froth flotation, as a preconcentration method, is a viable alternative to increase selectivity in downstream processes. In this work, the possibility of Alunite preconcentration with micro flotation and mechanical flotation cells was studied. The influence of type and dosage of collector, pH, solid percent, particles sizes were considered on the flotation. Flotation tests were designed based on Taguchi method, followed by performing based on surface charge differential between alunite and quartz on direct and revers flotation. In the reverse flotation, Ca+ 2 and Ba+ 2 were used to activate the quartz surface. Under optimum conditions of reverse flotation, alunite grade and recovery were 40% and 62.1% respectively. In the direct flotation, the surface of alunite was activated by F- . The grade and recovery were 50 and 89%, respectively.

Mar 2, 2022

By Nikhil Dhawan, SHAILA MIR

The emerging problem of electronic waste has promoted the interest in exploring metal recovery from secondary sources under the umbrella of urban mining. This study presents the physical and chemical characterization of various components present in end-of-life laptops to help ascertain the sites of metallic values and develop sustainable and environment-friendly recycling. The recycling potential of a laptop is evaluated by sequential disassembly, separation, and characterization of components such as body (49.8 wt.%), printed circuit board (9.7 wt.%), hard disk drive (4.9 wt.%), and battery (12.4 wt.%). The printed circuit boards comprise metallic values, majorly copper (25 wt.%), tin (5.8 wt.%), and lead (3.1 wt.%). The precious metals (Au, Ag) with an economic advantage are abundantly present in electronic components such as integrated circuits, capacitors, resistors, and processors. Critical elements such as lithium, cobalt, and rare earth elements are detected in the discarded laptops' batteries and hard disk drives. The recycling potential representing the amount of material that can be recycled in various components, depending on the type, concentration, and purity, is determined in the range of 36–100%. On average, a unit laptop contains~ 386 g of Cu (14.45 wt. %), 49.73 g Co (1.86 wt. %), ~346 mg Ag, ~141.2 mg Au, ~650 mg of rare earth elements (Nd, Dy). This study presents a database for investigating integrated and economical recycling techniques from different components of a laptop, and an overall process flow is also proposed.

Jul 24, 2021

By P. Eger, D. Green

"The Soudan Mine was Minnesota’s first iron mine, operating from 1882 to 1962. The mine was donated to the State of Minnesota in 1963 and became a state park in 1965. The mine discharge contains elevated copper and cobalt concentrations and has been treated with an ion-exchange-resin-based system since 2009. The system includes flow equalization tanks, bag and cartridge filters, a carbon tank and two commercial ion exchange tanks. Although effective, the system’s high cost, inefficient removal of suspended material and substantial maintenance have been ongoing and troublesome issues.In November 2012, we initiated a pilot test using a peat-based sorption media. This media is produced from raw reed-sedge peat through a patented, low-temperature carbonization process that produces a hardened granular sorption material. The granules are uniform, have high hydraulic conductivity and maintain the high metal affinity of natural peat.Mine water was pumped through the peat-based sorption media without any pretreatment. Copper input concentrations typically ranged from 30 to 60 µg/L but increased to a high of 364 µg/L in the summer of 2013. Since startup, more than 60 million L (16 million gal or 32,000 bed volumes) have been treated, with average removal rates of 75 percent for particulate copper and 60 percent for dissolved copper. Backwash is required at about 4,000 bed volumes, but with a combination of air sparging and high-flow backwash, the suspended material is removedfrom the media.One tank of the peat-based sorption media was found to remove copper similarly to the first four components of the existing system. By reducing size and complexity, reductions in operation and maintenance costs are possible. A comparison of the estimated costs indicate that replacing the four initial components with the single media tank has the potential to reduce annual operating costs by more than half from around $163,000 to $70,000."

Jan 1, 2015

By N. Arbiter

Introduction The theoretical analysis of tumbling mill energetics and performance has largely neglected mill dimensionality, and, in particular, the importance of the length/diameter (L/D) ratio. This is in spite of the fact that practice varies substantially: geographically, for semi-autogenous and autogenous mills, as between North America and South Africa/Scandinavia, and historically, for overflow ball mills, for which the L/D ratio has increased significantly from the earliest small mills to the largest mills currently. The present study is concerned primarily with the influence of the L/D ratio on the design and operation of overflow ball mills, on the occurrence of the overload phenomenon, and on the limits, if any, it may impose on mill capacities because of critical pulp axial velocity limits. It will be shown that the shape factor is of major importance in this area and that its adjustment to the extent that this is practical should remove the diameter limitations previously postulated for this mill type. Dimensionality in mill design Ball mill shape factors in the period prior to 1927 (Taggart, 1927) averaged 1.1/1 for 29 center discharge mills and 1.0/1 for 30 peripheral discharge mills. With the resumption of new plant construction after the 1930s depression, the Morenci concentrator continued the 1/1 ratio with its 3.1 x 3.1 m (10 x 10 ft) mills. The ratio was increased progressively from then on, reaching 1.6 and 1.8/1 for the largest overflow mills currently. As shown recently (Arbiter, 1989), this is in sharp contrast to autogenous and SAG mill shapes, for which the ratio averages 0.4 in North America. On the other hand, South African practice, starting at the turn of the century with autogenous mills having 4/1 ratios, moved toward 1/1 until recently, when a 2.5/ 1 ratio mill was installed. The reasons given for such divergent practice for mill shape factors are in some respects contradictory and generally inconclusive. The most complete discussion from a practical viewpoint (Dor and Bassarear, 1982) is limited to primary SAG and autogenous mills. Considerations of ball mill dimensionality have had a twofold direction. On the one hand, it has been argued that ball mill efficiencies should increase with increasing diameter and that the specific energy for a particular grind should be reduced accordingly. An inadvertent test of this idea at the Bougainville operation (Burns and Erskine, 1983) resulted in drastic underpowering, which led to failure to reach design capacity until additional mills were installed. This can be taken as strong evidence against any increase in efficiency with diameter. In another direction, it has been argued (Arbiter and Harris, 1982, 1983) that there is a limit to ball mill diameters because of the demonstrable limit to axial flow velocities evident in the overload phenomenon, which as a fact is incontrovertible. But that it places a limit on mill diameters overlooks the evidence given below that appropriate variation in the L/D ratio will permit major increases in diameter, limited only by constructional or economic factors. The present study was directed toward quantifying the overload phenomenon through examination of the influence of mill dimensionality variation and mill operating variables on its occurrence. It is shown that varying the operating conditions, specifically the load fraction and the fraction critical speed, can reduce the risk of overload for existing operations; while appropriate decreases in the L/D ratio can minimize the risk in the design of new circuits. Ball mill overload Ball mill overload is a consequence of the approach to a critical velocity with increasing feed rates or circulating loads. Although the effect has been known for over 50 years, there have been no previous attempts to quantify it. The following description of the ball mill as a flow system is the preliminary to a quantitative analysis: 1) In the absence of a ball load, axial flow of pulp through a mill resembles open channel flow, except for disturbances near the shell due to shell/lifter rotation. 2) For a given ball load (Lf), the void fraction available for pulp hold-up (H) for the ascending portion of the load is approximately 0.4 Lf. In the descending portion, it is greater than this and increases with fraction critical speed (Fc) because of load expansion. 3) Increasing the feed rate increases pulp hold-up and progressively fills the available void space. At a critical flow rate, which depends on system geometry, hydraulic head and pulp rheology, void filling reaches its limit; a pool forms rapidly and fills available space outside the ball load and up to the overflow level. Prior to this, pulp discharges mainly along the ascending rim of the overflow. 4) For a small mill (Lo et al., 1990), it has been shown that with increasing feed rates the critical filling is at or near 50% of the mill volume. Power drops rapidly when this level is reached, as required by the torque formula. 5) The transition to overload is associated with the following phenomena: a) The decrease in power draw. b) Damping of mill sound. c) Reduced comminution of coarser feed sizes, probably due to reduced direct impact in the presence of a pool. d) Increased circulating loads, which further intensify the overload, and increased density of cyclone underflows, which can lead to roping. e) Conditions beyond overload are not known because feed rates are not delibelrately increased beyond this point. f) The existence of the phenomenon limits the capacity of a mill with a fixed set of operating conditions and can prevent the balancing of hard ore feed rate decreases by increases in soft ore rates.

Jan 1, 1992

By I. R. Miller

The purpose of this discussion is to try to offer some suggestions as to how the zinc and lead markets will perform over the balance of the decade, review some of the factors influencing production and come to some conclusions regarding prices. The past 2 years have been simply awful in terms of consumption, prices and overall profitability for the industry. In looking ahead to the year 2,000, it is probably not a bad idea to look back about 20 years to see where we have been and how we got to where we are. Zinc consumption from 1950 to the early 1970's increased steadily and rather nicely at an annual average rate of close to 4.0%. It was starting in the early 1970's that things started to get volatile and at times, difficult for the industry. The general economic boom of the early 70's was marked by annual consumption rates close to 5.0 million tons, a level not seen again for over 10 years. The First Oil shock of 1973-74 was followed by the sharpest consumption drop in memory. In two years, Western World zinc consumption fell 27%. While our chart shows a rather dramatic recovery thereafter, the average annual growth rate since 1970 has only been a rather modest 1.5%.

Jan 1, 1992

By Jr. Koch

In the Hydrogeochemical and Stream Sediment Reconnaissance (HSSR) phase of its National Uranium Resource Evaluation (NURE) pro¬gram, the U.S. Department of Energy (DOE) samples about 1400 sites for stream sediments in each two-degree quadrangle of the National Topographic Nap Series (NTMS) and about 1200 sites for ground or stream water. Through a geological/statistical analysis, we find that NURE data are useful for assessing the uranium resources of the NTMS quadrangles. Our computer-implemented System for Uranium Resource Evaluation (SURE) calculates a score for each NTMS quadrangle based on the HSSR, geologic, and radiometric data; we score the four Colorado quadrangles--Pueblo, Montrose, Durango, and Trinidad--that comprise the study area. We interpret the scores in terms of geology and the known occurrences, which include the Hanson ore deposit with reserves of 30,000,000 lbs. U308 and the Pitch mine with reserves of 1,740,000 lbs. SURE depends in part on distinguishing mineralized or metalliferous granitoids from barren ones through "geochemical signatures."

Jan 1, 1980

By R. A. Ynchausti

In January 1993, Pyramid Resources began implementation of a supervisory expert control system at a U.S Cu concentrator. The objective of this control system was to improve plant profitability through unit operation optimization and integration. This paper documents the key aspects of the project, including: economic and associated benefits, system architecture, concepts of implementation, process control strategies, practical aspects of implementation and conclusions. The implementation of expert control at this concentrator has resulted in an estimated benefit of $50,500 per day. This is a net result of improved tonnage, reduced reagent consumption, improved concentrate grade and reduced energy per ton produced.

Jan 1, 1995

By B. K. Slone

The depth at which it becomes more economical to use slope access as opposed to box-cut access for near-surface coal reserves, termed "transition depth," can be established for a given coal property by considering cost estimates for those parameters that determine respective access method. Fourteen parameters that influence access cost were identified and ranges determined. A Transition Depth Formula was developed to permit determination of the better method of accessing a reserve, given such information as highwall slope, road width, total bench width, pit dimensions, road grade, belt component costs, pumping costs, and cost of overburden removal and reclamation. In most cases, in Appalachia, studies indicate that the transition depth ranges between 20 and 30 m.

Jan 1, 1998

By C. Beul, R. Schipper, L. Santisteban

The Garfield Mine is a legacy vanadium mine in western Colorado with multiple mine openings, waste rock piles, and adit seepages that Cyprus Climax Metals Company (a Freeport-McMoRan Inc. subsidiary) voluntarily remediated through the Colorado Voluntary Cleanup Program (VCUP). The process to obtain a No Further Action determination from the VCUP is presented. Water treatment of the adit discharge, which is not part of the VCUP, is also presented. The remedial action objectives (RAOs) included preventing direct human or biotic exposure to the waste rock and radiation emitting from the waste rock; reducing metal concentrations and radionuclide activities in adit discharge and keeping the discharge from leaving the site; maintaining the existing undeveloped character of the surrounding landscape; and eliminating the need for a stormwater permit. The design included regrading, installation of an infiltration barrier and rock cover, a diversion channel, and a biologically based passive remediation water treatment system. An initial lab-scale evaluation provided design criteria for the full-scale system, which consists of a sulfate-reducing biochemical reactor and post-treatment aerobic polishing cells. The system has been operational since construction was completed in late 2016. Results indicate that the treatment objective of 90% bulk removal of constituents of potential concern (COPCs) is largely being met. The system also provides year-round operation and non-point discharge of effluent. After 20 plus years of site investigations, the remedial action was completed in 2016 with all RAOs met and a No Further Action determination granted in 2018. INTRODUCTION The Garfield Mine is a legacy vanadium mine located approximately 11 miles north-northeast of Rifle, in Garfield County, Colorado. It was leased and operated intermittently from 1929 to 1975 but was never operated by Cyprus Climax Metals Company (Cyprus) (a Freeport-McMoRan Inc. subsidiary). Prior to reclamation, site features included: • six mine openings—five adits and a vertical opening; • two adits discharging mining-influenced water (MIW)— Adit 1 at less than 4 liters per minute (L/min) and Adit 2 at 7 to 10 L/min; • main waste rock pile (13,500 m3) and small waste rock pile (minimal volume); • debris (small structure, mattress, bottles, shells, etc.) indicating site used for recreational purposes; and, • unimproved road, which has been used to access forest fires.

Jan 1, 2019

By W. D. Eddy

Topaz is a fluorine-bearing aluminosilicate mineral of the kyanite-sillimanite group. The formula has been given as Al2SiO4(F,OH)2, [Al(F, OH)]2SiO4, or 2Al2O3?2Al(F, OH)3SiO2. Topaz is generally considered a semiprecious-to-precious gem mineral, but it is a fairly common minor constituent of granites. In 1937 the U.s. Geological Survey reported an occurrence of massive topaz in the 3 Brewer gold mine, Chesterfield County, S.C. Over the next 8 years considerable research was done on topaz to establish its properties and to determine possible end uses The first reported shipment of ore was in 1939 Peak production was reached in 194412, and output then began declining until the mine closed in 194913? During World War II topaz was used in refractories, as a flux in steel mills, and in the manufacture of borosilicate glass 14. The main advantage of using topaz over kyanite and sillimanite in refractory use was that calcined topaz converted to over 95 percent mullite (3A1203 .2Si02) while calcined kyanite or sillimanite contained only 88 percent mullite15. However, a major draw-back to utilization of topaz is the evolution of fluorides during calcination. In 1973 the Geological Survey reported that topaz occurred in probable commercial quantities in at least two domestic locationsl6: the Brewer area of South Carolina and a location in Colorado. The literature also indicated that progress had been made in controlling fluorine-bearing fumes. This research has progressed to the state where valuable fluorine compounds are being recovered on a commercial scale 17. Thus, a major obstacle to utilization of topaz appears to have been resolved. The Bureau of Mines under its program of advancing minerals technology initiated a project to study techniques for the concentration of topaz from disseminated materials. The objective of the study was to determine those techniques that would be used to produce topaz concentrates that could be utilized as a source of fluorine and mullite.

Jan 1, 1977

By Sheshang Singh Chandel, Prvan Kumar Katiyar, Navneet Singh Randhawa, Prince Kumar Singh

Mineral process wastes, such as metallurgical slags nowadays, are of significant metals and materials resources in the circular economy. The usability of iron-making slag is well established; however, steel slag still needs to be utilized due to various physicochemical constraints. In this context, we have reviewed the global steel slag generation, processing, characteristics, and potential application with suggestions for futuristic possibilities to address the technical, economic, and environmental aspects of steel slag recycling. The steel slags inherit characteristics from ore chemistry and their processing routes, which may be Blast Furnace-Basic Oxygen Furnace (BF-BOF) or Direct Reduced Iron-Electric Arc Furnace (DRI-EAF) steel-making routes with/without treatment in a Ladle Furnace (LF). The average generation of slags in BOF, EAF, and LF processes accounts for 110 kg, 70 kg, and 40 kg per tonne of steel, respectively. Considering global steel production of 1878 million tonnes in 2020, the associated slag by-products possess a significant resource of major components like FeO, CaO, SiO2, Al2O3, etc. Selecting a recycling technique depends on the slag's end use, logistics, and physicochemical characteristics. BOF and EAF slag are used ~ 48% in road construction applications and ~ 10% for industrial specific metallurgical use. EAF slag is preferred as cementitious material as compared to BOF. Hot-stage modification of BOF slag has emerged as a preferred method to achieve desired chemical and subsequent mechanical properties for end use. BOF slag has high carbon dioxide sequestration potential for in-plant remediation among steel slags.

Nov 20, 2023

By S. Komar Kawatra, J. J. Carlson

"It is widely known that steel made from iron ore pellets high in phosphorus is brittle. Removal of phosphorus from iron ores such as hematite has proven to be difficult. The majority of research on phosphorus removal from hematite ores has been focused on flotation. The purpose of this study was to assess phosphorus distribution under current operations at an existing hematite concentrator and determine how efficient the plant is at removing phosphorus. Objectives included: (1) Perform a phosphorus balance at an operating plant to assess phosphorus distribution under current operations; (2) Determine phosphorus removal efficiency of each separating unit and (3) Perform a quantitative liberation analysis to determine max phosphorus removal by physical means. Liberation was quantitatively determined using wet high intensity magnetic separation (WHIMS) to separate iron ore from liberated apatite (and other nonmagnetic compounds).It was observed that 61.2% of the total phosphorus entering Plant F was rejected during current operations. The desliming process was the most efficient step at removing phosphorus (45.1% P removed), whereas the flotation step was much less effective (16.1% P removed). Liberation analyses indicated that the deslime thickener may have been more efficient at removing phosphorus than the flotation circuit due to the greater abundance of liberated phosphorus in the deslime thickener feed. The scavenger returns stream was identified as having a high concentration of phosphorus and a possible point of phosphorus removal. However, liberation analyses indicated that only 32.4% of the total phosphorus in the stream was liberated. Grinding the scavenger returns to 80% passing 5.2 µm increased the amount of liberated phosphorus in the stream to 70.4%."

Jan 1, 2014

By Albert Hofstra, Joel Leventhal

The gray, dark gray and black colors of the sediments and the presence of pyrite in the Carlin, Jerritt Canyon, Horse Canyon, Betze, and Gold Acres sediment-hosted disseminated gold (SHDG) deposits indicate that these rocks are not oxidized with respect to carbon and iron sulfide. Both ore and non-ore host rocks generally are dark colored. Analyses for noncarbonate carbon (usually called organic carbon) in samples from SHDG deposits and related unmineralized rocks indicate that these rocks contain from a few tenths to a few percent carbon. We use the term noncarbonate carbon because, strictly speaking, organic carbon denotes carbon that is bound to hydrogen. However, the carbon present in SHDG deposits and related rocks has very little hydrogen chemically bound to it. Rocks with less than 2% organic carbon are usually gray in color. The dark colors in our samples are due to the relatively high temperatures that these rocks have experienced relative to organic matter maturation. We confirm the results of Nelson et al. (1982) that previous reports (for example, Radtke and Scheiner, 1970) were incorrect in reporting significant amounts of humic acids, extractable hydrocarbons (bitumen) or "carbonaceous" material. We believe these reports had problems with contamination or misdirected emphasis. Our samples consisted of unweathered rocks, collected mainly from mine cuts (benches, trenches, cores) and put into glass containers or wrapped in aluminum foil to avoid contamination. Only a few samples were collected from natural outcrops. We were careful to avoid samples in proximity to blast holes that were packed with ammonium nitrate and diesel fuel ("ANFO"). Contamination by ANFO may explain some of the previous reports of hydrocarbons, pyrobitumens, humic acids and anomalous nitrogen. When we were collecting samples, it was apparent that in addition to lithologic and color differences, the surface appearance of the rocks was distinguishable: "preg robbing" material that had a sooty surface; "graphitic" rock with a shiny surface; and more ordinary rocks that had a flat gray or black appearance. [Preg robbing is a term for ores that adsorb gold from the milling circuit; the graphitic material had not been x-rayed to actually identify graphite.] We hoped that laboratory study of these samples would show some fundamental differences that we could relate to ore genesis. Even though these sample types appeared different in the field, we could not distinguish any differences in thermal maturity (see below). Normal petroleum/organic geochemical laboratory procedures (solvent extraction, Rock-eval, pyrolysis-gas chromatography) failed to yield much information because the carbon was over-mature relative to oil and gas generation. Rock-eval results show less than 20 (± 20) mg of pyrolizable hydrocarbons per gram organic carbon (and an indeterminate Tmax) indicating that the organic matter in these deposits is burned-out relative to petroleum generation [with an H/C (atomic) ratio of less than 0.41. Because of the non-definitive results from standard organic geochemical methods, x-ray diffraction (XRD) and Raman spectroscopy were used to characterize the carbon. Prior to XRD, samples were treated with HCl and HF acids (1) to remove quartz that has a peak near 26° [20] at the position of the 002 peak of graphite and (2) to concentrate the carbon to a level detectable by XRD. The residue was put on a colodion mount and analyzed by XRD. The XRD analyses yielded a broad peak at 26° [20]. The shape of this peak indicates that the carbon is more thermally mature than anthracite rank coal but less mature than crystallized graphite. Based on XRD results from the literature (Landis, 1971) that are calibrated to metamorphic facies, results from our samples correspond to pumpellyite-actinolite to lowermost greenschist facies conditions, probably

Jan 1, 1990

By D. H. White

The term "block caving" as used in this text includes panel caving and mass caving as described by Julin and Tobie (1973). This method reportedly was conceived in 1895 when a block of ore ranging up to 200 ft (61 m) in width was undercut at the Pewabic Mine on the Menominee Range in Michigan (Bucky, 1956). Over the years this method has evolved largely as a consequence of trial and error in which operators became acutely aware of what nature would allow and what it would not. Today block caving remains somewhat of an art rather than a science where experience often dictates what needs to be done. With the plunging prices of base and specialty metals in the early 1980s caving mines sought to reduce operating and development costs in order to return to profitability. For some this became an impossible task given the design, grade and operating costs of the day and closure became the only alternative. One lesson learned from the experiences of the 80s is that survivability depends on cost competitiveness and this implies the efficient use of resources and technology. To this end applied rock mechanics serves as a useful tool. The purpose of this paper is to show how Geomechanics can be used? in block caving to facilitate sound decision making and how this equates to dollars saved. Features of mining method selection, design and operation are discussed as related to this theme.

Jan 1, 1990