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By L. Wilfley, G. A. Calboreanu

"Wear resistant 25% Cr hard cast iron is a heterogeneous material containing 1,600 HV hard chromium carbides embedded into martensitic matrix. It is the most used cast alloy in wear operations of mining industry, chemical industry, fracking, etc. During the last decade the hard iron manufacturers increased wear resistance in the following ways: • Reduce / eliminate harmful retained austenite and chromium + carbon depleted areas especially within eutectic carbide cells. In Max5A the improved 25% Cr alloy brittle carbide eutectic cells are removed which increases toughness. • Delay delamination between chromium carbides and matrix particularly during sliding abrasion wear. Delamination commences at the weakest chromium carbide – matrix contact lines exposing chromium carbide to localized bending stresses which gradually chip carbides away. Once the chromium carbide base is removed accelerated wear degradation occurs. • Evenly randomize and avoid discontinuities in the chromium carbide network. When chromium carbides are perpendicularly positioned to the wear surface wear rate is minimized. On the contrary the highest wear rate takes place when 3D oriented chromium carbides are parallel with the wear surface. In this case chromium carbides cover top network discontinuities being just matrix supported which leads to rapid carbide cracking. The new hard iron discussed in this paper fulfils and exceeds the above requirements increasing the casting service life around 250% in various wear operations. Physical and mechanical properties of Max5A are compared with well manufactured Industry Standard 25% Cr hard iron. The paper presents practical accomplishments obtained in the USA, South America and India. INTRODUCTION The chemical composition of Max5A and Industry Standard 25% Cr are identical like specified by ASTM A532 Class III Type A requesting C = 2.0-3.3% and Cr = 23.0-30.0%. Manganese, silicon, nickel, molybdenum and copper contents are limited by maximum accepted quantities.[1] According to published data the eutectic chromium carbides chemical composition of 25% Cr hard iron is Cr = 54% and C = 8% which indicates strong depletion of chromium and carbon along chromium carbide - matrix interfaces. [2] The optimized Max5A microstructure has minimized retained austenite, is devoid of harmful carbide eutectic cells and has the most advantageous chromium carbide and secondary chromium carbide morphology. This favorable microstructure increases wear resistance, mechanical properties and corrosion resistance. The paper compares the Max5A microscopic features and mechanical – physical properties with similar properties of high quality Industry Standard 25% Cr."

Jan 1, 2017

By W. C. Lacy

Traditionally Australia has been a leader in underground mining methods with world standards being established at Mount Isa, Broken Hill, Kalgoorlie, and Mount Lyell. With decline in higher-grade reserves and discovery of a multitude of near-surface oxide gold de-posits attention has shifted to bulk mining of low-grade gold ores by open pit methods, heap leaching and activated carbon recovery of gold values. The paper examines historical change in mining philosophy and the economic effects of isolation.

Jan 1, 1993

By José Maria Silva, Eugênio Oliveira, Ruberlan Silva, Thiago Souza, Lucas Reis, Lorena Guimarães, Matheus Bianchetti

Copper mine operations consistently generate a large quantity of tailings after their concentration processing. These tailings are either piled up or stored in a dam, where both alternatives of disposal lead to environmental impacts. The dam copper tailings have a huge potential for by-product production thus improving sustainability and profitability in industrial operations. This paper describes a study aiming at producing a magnetite pellet feed as a by-product from a dam chalcopyrite flotation tailings and generating free magnetite non-magnetic rougher tailings. The chemical assay and mineralogy analysis of dam copper flotation tailings indicate low Fe grade (6.34 wt%), high SiO2 grade (63.14 wt%), low magnetite (3.6 wt%), and high iron silicates content (22.72 wt%). The complex mineralogy of these tailings leads to the need of grinding the rougher magnetite pellet feed as fine as 80% of particles smaller than 20 μm ( P80) aiming at improving magnetite liberation. The proposed process route comprises a scalping screen, low field rougher, and cleaner magnetic separations using a drum magnetic separator (80 kA/m or 1000 Gauss), followed by silicates reverse cationic flotation as a recleaner step. With this process route, it was possible to produce a high-quality magnetite pellet feed reaching 68.00 wt% of Fe, 2.13 wt% of SiO2, 0.51 wt% of Al2O3, and 95.1 wt% of magnetite. The Blaine specific surface area was 2300 cm2/ g, being in the range recommended for pelletizing (1900 up to 3000 cm2/ g). The total mass, Fe, SiO2, Al2O3, and magnetite recoveries were 2.77%, 29.71%, 0.09%, 0.12%, and 73.17%, respectively. Most of the magnetite losses, around 20% of total magnetite fed, occurred in the oversize of the scalping screen. Additionally, it generated non-magnetic rougher tailings with less than 1.0 wt% of magnetite that could be used as by-products and reduced the amount of tailings disposal in a dam. This study contributes toward the sustainability of operation, minimizing tailings disposal.

Sep 28, 2022

By R. M. Rowe, T. E. Irwin

A flowsheet was developed for upgrading Climax tin concentrate to about 45 percent Sn. Bench-scale test work conducted at AMAX Extractive Research and Development, Inc., was used to design a multiple-stage tin- processing flowsheet which included screening, high-intensity magnetic separation, and high- tension (electric- static) separation. Formal engineering and plant installation was accomplished in-house by Climax mill personnel. The successful start-up operation of the facility was the joint effort of the AMAX R&D and Climax-mill staffs.

Jan 1, 1986

By Terry I. Mudder

Metal complexed cyanides in wastewaters form as a result of interactions of free cyanide with metals present in the wastewater and exhibit varying degrees of stability, toxicity, and treatability. Thiocyanate, a pollutant commonly found in cyanide bearing wastewaters, is formed through the interaction of free cyanide with sulfur containing species (i. e. pyrrhotite) both present in the wastewater. In certain industrial processes, such as the beneficiation of gold and silver, cyanide is an essential reagent. Since free cyanide, complexed cyanides, and thiocyanate are potentially toxic to humans and aquatic organisms, these compounds must be removed from wastewaters prior to their discharge into surface or ground waters serving as potential potable water sources, marine or fresh water habitats

Jan 1, 1984

By L. G. Austin

The model originally proposed by Austin, et al., has been modified in the light of further investigation of the breakage laws in SAG milling. The model uses a fully-mixed reactor for short L/D mills and calculates the breakage of small particles by the nipping action of balls and pebbles. The autogenous self-breakage of larger lumps is treated as an effective means of fast breaking rock and much slower breaking rock. Experiments show that fast autogenous breakage can be treated as a first-order breakage with an overall specific breakage rate which incorporates fracture and chipping. Similarly, slow breakage incorporates fracture and abrasion. Autogenous breakage is cushioned by the accumulation of fine material. Simulation results are compared to results from tests in a 2 m diameter continuous pilot-scale SAG mill. Future work is discussed.

Jan 1, 1985

By V. K. Gupta

The experimental results of some wet batch and continuous grinding experiments carried out in a 40 x 40 cm grate discharge ball mill have been analyzed for the variation of the grinding rate parameters, Si, with various operating conditions using a linearized population balance model. The wet grinding tests were carried out under conditions which were very similar to those encountered in the actual practice in plant operation. Well distributed -4 and -20 mesh feeds of two complex pyritic ores were used. The volume-percent solids in the pulp ranged from 30-40, the particle load from 5-8 kg, and the solids feed rate from 70-1 10 kg/hr.

Jan 1, 1981

By Ross W. Whipple

Applying geophysical techniques in potential geothermal areas has received: little attention in the United States, though it is widely used and accepted in New Zealand and Italy. In those countries it has been demonstrated that geophysical surveys can be useful in locating new geothermal areas and in directing test drilling programs. Geophysical studies in a thermal region can provide structural information and, if conditions are right directly; locate subsurface hot waters. They can give indications of sub-alluvial rock types and of the thickness of alluvial cover. In certain cases they can be used to estimate the power potential of a hydrothermal system.

Jan 1, 1963

By Aditya Rana, A. K. Mishra, M. M. Singh, N. K. Bhagat

Many cases of derailments, train accidents and traffic interruptions due to the slope failure and rockfall have been reported in the past on Konkan Railway trackline because of unstable rock slopes. The authorities installed stability enhancement measures such as wire netting, rock bolting and shotcreting which were found insufficient. The failed geotechnical measures prompted authorities to excavate the unstable slopes. Rock blasting is generally not preferred for the excavation of slopes along the transportation route. Stabilisation and widening of slopes by blasting need closure of the route for a considerable time. Discontinuing traffic on Konkan Railway route was undesirable as it connects two important port cities Mumbai and Mangalore. The paper presents an innovative directional controlled blasting (DCB) technique for excavating three unstable slopes (i.e. cuttings) in hilly terrain without disturbing the traffic. New empirical formulae for small geometry blasting to estimate burden and throw have been developed. A unique sequence of firing and excavation with powder factor varying between 0.04 and 0.39 kg/m3 irrespective of the actual free face was used for rock blasting. The process controlled the throw of blasted material over the track and resulted in a smooth and stable wall. The slope angle has been changed from 80 to 82° to 45–47° and probability of wedge, planar and toppling failures was eliminated. About 100 such accident-prone slopes (52 km cumulative length, 7% of total route) situated at discrete locations along Konkan Railway route were stabilised using the technique. No incident of rockfall and slope failures along these slopes has been observed after the stabilisation.

By J. A. Hartlage

Four years ago, Ashland Chemical Company introduced a new copper solvent extraction reagent for hydrometallurgical utilization. The reagent, an alkylated 8-hydroxyquinoline, is called Kelex 100 ®(1). Since the time of introduction, the reagent has been investigated by several workers for its solvent extraction utility in the separation and purification of transition metals (2-7). In the following text, those factors which are essential to the optimum performance of Kelex 100 on concentrated copper solutions will be discussed. To date, the reagent has been found useful in active concentrations of from 1 to 20% Kelex 100. In order to achieve the best performance for Kelex 100, it has been found that a phase modifier is essential. Various alcohols and phenols have been evaluated in order to improve phase performance and stripping characteristics of the solvent system. Modifiers and diluents are especially important when solvent activities between 3 and 20% Kelex 100 are employed. A 15% Kelex 100 solvent, for example, containing 15% nonyl phenol in Escaid 100 (Exxon) can be utilized to remove greater than 33 grams per liter of copper from an aqueous feed without interstage neutralization, and in the same process scheme the reagent can be stripped to less than 1 gram per liter barren organic producing a pregnant electrolyte containing greater than 40 grams per liter copper and less than 120 grams per liter sulphuric acid. The unusually efficient stripping of Kelex is accomplished by protenation of the extractive molecule during the stripping process.

Jan 1, 1974

By J. O. Marsden, M. M. Botz

"Heap leaching is responsible for approximately 21 percent — 3.9 million metric tonnes — of copper production and 9 percent — 270 metric tonnes or 8.7 million ounces — of gold production worldwide. Given metal price assumptions of $2.25/lb for copper and $1,250/oz for gold, these portions of global production generate revenues of $19 billion and $11 billion, respectively. For heap leaching operations, small changes in metal extraction efficiency drive substantial changes in operating cash flow margins and affect project viability. Variances in actual metal production from budgetary forecast estimates can mean the difference between success (value creation) and failure (value destruction) for a heap leach project. Despite the importance of heap leaching for copper, gold and silver production worldwide, there are no clear standards or guidelines for developing heap leach metal production forecasts. There is considerable variation among copper and gold producers that apply heap leaching technology as to how such modeling and forecasting is applied at the operational level. There are a variety of different modeling approaches that can be utilized for this purpose. However, these vary in cost, complexity, time to implement, accuracy and credibility. This paper explores the options available for heap leach modeling, discusses the advantages and disadvantages of each approach, and provides guidelines/recommendations for an effective approach to modeling metal production at existing or planned heap leach operations. The paper also provides expected ranges of variation between actual and modeled production.Introduction and backgroundHeap leaching is fundamentally different from other methods used in the processing and extraction of gold, silver, and copper and other base metals. Considered globally, the technique represents a relatively modest portion of production compared with the primary technologies applied for copper and gold recovery. Approximately 21 percent of copper production is attributable to heap leaching techniques, as opposed to 79 percent by conventional crushing, milling, flotation, smelting and refining. For gold, about 9 percent of global production is obtained from heap leaching, compared with over 90 percent by various combinations of crushing, milling, gravity concentration, flotation and/or agitated cyanide leaching processes or, in the case of placer deposits, by direct gravity concentration and/or amalgamation. In cases where heap leaching is the primary source of metal production, the ability to accurately forecast metal production may be the difference between commercial success and failure for a project or operation. Heap leaching — of crushed or run-of-mine ore — is also commonly applied as a secondary process to treat lower-grade ores from a given deposit, such as at Chuquicamata in Chile, Cerro Verde in Peru and Bagdad in Arizona for copper extraction and at Twin Creeks and Carlin, both in Nevada, for gold. In such cases, heap leaching provides incremental revenue to supplement the primary process. Variances between actual and forecast metal production from heap leaching as a secondary process are typically tolerated and absorbed within the overall site metallurgical balance. In either case, accurate production modeling and forecasting creates significant value for heap leaching projects, is essential for managing production expectations — timing and volumes — and is likely to be a critical success factor for most heap leach projects."

Jan 1, 2017

By W. R. Reed, G. W. Luxbacher

Mine worker exposure to respirable coal dust and silica dust can result in coal workers’ pneumoconiosis and silicosis which are occupational respiratory diseases that have no cure and are ultimately fatal. In 2021 the Mine Safety and Health Administration reported just over 11,000 operating metal/nonmetal mines with the potential for silica dust exposure for just under 200,000 miners. The National Institute for Occupational Safety and Health (NIOSH) conducts respirable coal and silica dust research associated with mining through its Office of Mine Safety and Health Research. NIOSH respirable coal and silica dust research began under the U.S. Bureau of Mines (USBM), which initially examined silica dust, progressing to include coal dust, with a stronger focus after 1969 with the passage of the Federal Coal Mine Health and Safety Act. A review of extramural research to develop real-time respirable silica dust monitor is presented along with a brief history of silica dust regulation.

Feb 1, 2024

By Nicholas Tabor, Norbert Fuegenschuh, Richard Taylor, Jean-Marc Wehrli

The Plymouth Tunnel forms part of the 16-mile long route of the Purple Line Light Rail Project promoted by the Maryland Transit Administration. The tunnel is approximately 1,220 feet in length with a shallow vertical alignment located beneath a residential area. Challenging mixed ground conditions included surficial soils, decomposed Saprolite, and highly weathered to fresh rock requiring the use of the Sequential Excavation Method (SEM) for construction. This paper discusses innovative approaches implemented during construction as well as challenges faced during planning and installation of a spray-applied permanent concrete liner.

Jun 13, 2021

By F. A. Edwards

The Macassa Division of Lac Minerals Ltd. is a high grade gold mine that has been operating in Kirkland Lake, Ontario for the past 51 years. It produces approximately 120,000 tonnes of ore per year, averaging approximately 18.7 grams per tonne (0.45 ounces per ton) of gold. Over the years, the ore body has been developed to a point where mining takes place below the 1,220 m (4,000 feet) elevation and 1,830 m elevation (6,000 feet) west of the original shaft. The mine is currently serviced by two surface and two winze shafts. (Figure 1). The ore body extends below the 1,930 m (6,400 feet) Level which is the bottom of the existing mining operation. An ambitious underground exploration program has substantially increased these reserves. Concurrent with the discovery of greater reserves, has been rapidly escalating mining costs, particularly labour which constituted 66% of the total cost in 1981. These costs continued to escalate as the major proportion of mining was carried out to the west, far away from the existing hoisting systems. This required a three lift hoisting system for men, materials and ore. In addition to the increased travel time and ore handling expense, increased humidity, temperature, ventilation and compressed air losses have added to the inefficiency of the operation. As a result, studies were initiated in 1980 to assess the future of the Macassa operation. The results of the studies indicated that a new deep shaft with a potential to hoist from 2,135 m (7,000 feet) or below was the most attractive option. Detailed design of the shaft and associated surface plant commenced in January, 1982 and the following milestone dates were achieved: 1980 - initial study of hoisting and haulage facilities; January, 1981 - ore reserves in excess of 1 million tonnes of ore at 18.7 g/tonne (0.45 oz/ton) confirmed; March, 1982 - preliminary study of Macassa No. 3 Shaft completed; April, 1982 - permanent hoisting plant purchased; June, 1982 - site clearing commenced; September, 1982 - shaft collared; May 1, 1983 - shaft sinking commenced; February 11, 1985 - shaft reached 1,600 m (1mile); December 1, 1985 - shaft scheduled for completion.

Jan 1, 1985

By Gordon R. Haworth

Conditions affecting minerals investments are continuing to change at an increasing rate, and in this paper we will attempt to predict the course of mine financing in the future. We will first review the situation historically and then examine the problems and constraints that are now making the financing task more difficult. Suggested solutions are discussed, and a new method of financing a group of mining projects is hypothesized.

Jan 1, 1977

By Lloyd M. English

Accurate knowledge of the gas content of coal seams has always been important, according to the work of W. P. Diamond and J. R. Levine, 1981, but is becoming increasingly important for several reasons: Excessive gas emissions in the mine can create explosion hazards. Depletion of shallower coal reserves is forcing underground operations to greater depths where coal gas content is often greater. Natural gas prices and the need for new energy sources make naturally occurring coal gas a potential resource of economic interest. Information on coal gas content will impact mine designs, mine ventilation system specifications, and the effectiveness of methane control procedures used. Accurate resource assessment is needed for projecting and designing potential gas utilization enterprises. The emphasis upon methane hazards in mining operations, and upon potential methane utilization, has led us to improve upon the procedures used to determine coalbed methane levels.

Jan 1, 1984

By R. J. Roman

The results of column leaching tests on 3 different mixed oxide/sulfide ores from a Chilean deposit are presented. The 3 samples contained different ratios of oxide copper minerals to chalcocite and chalcopyrite. The results of the column leach tests are interpreted in light of the shrinking core model for coarse particle leaching. The copper recovery from the oxide minerals and the sulfide minerals each conforms to the shrinking core model with the total recovery equal to the weighted average of the individual minerals. The use of column test results to predict field heap leaching results using the multi-mineral shrinking core model is discussed.

Jan 1, 1993

By J. A. Barsin

Stationary air pollutant sources established prior to the New Source Performance Standard (NSPS) are not, at present, subject to federal emissions regulations. However, proposed legislation could require many existing stationary sources to reduce emissions. This paper will report upon the laboratory work undertaken to date by B&W to commercialize a limestone system for sulfur capture both in pulverized coal and cyclone coal-fired utility boilers. Updates will be provided on the US EPA/State of Ohio/Ohio Edison/B&W sponsored pulverized coal full-scale limestone injection demonstration, and on the EPRI/Baltimore Gas 8 Electric, Atlantic Electric/B&W sponsored laboratory-scale cyclone coal limestone injection investigations.

Jan 1, 1986

By Arthur W. Ruff

The iron-ore beneficiation and shiploading facilities of the Marcona Mining Company are at San Nicolas Bay which is on the Peruvian coastal desert about two hundred and seventy miles south of Lima. The mines are at elevation of 2,500 ft. Crude ore is transported from the Mine primary crushers to the beneficiation plants by ten miles of overland belt conveyors. After the predominantly magnetic ore is reduced to minus 3/4 in. in the secondary crushing plant, the ore is ground by rod and ball mills and concentrated with magnetic separators. The concentrate slurry is pumped to a thickening and filtering installation for desliming and dewatering. Pelletizing is accomplished with a Lurgi and also a Dravo-Lurgi machine. Deep-water shiploading facilities are available close to the beneficiation plants. Figure 1 shows a flowsheet of the general operation. Pelletizing operations are scheduled 24 hours a day and 365 days per year. A maximum tonnage rate with a high percentage of operating time is necessary to produce the three and one half million tons per year required from the two pellet lines. Process control is a major responsibility and of primary concern to all Operating personnel. The particle size, moisture, and chemical composition of the filter cake must be closely controlled to get uniform high-grade pellets. Key points are sampled routinely and analyzed on a two-hour basis to provide the necessary structure and chemical data to maintain a superior final product. The iron content of the pellets is held between 65.3 and 65.8 percent with 4.0 percent Si02. Sulfur is below 0.04 percent and copper is below 0.02 percent in the finished pellet.

Jan 1, 1968

By Kanhaiya Mishra, Phanil. K. Mandal, S. K. Behera, P. S. Paul, Prashant Singh, C. N. Ghosh

The techno-economic and social benefits of cemented paste backfill (CPB) resulted in its wide acceptance by the mining industry. The Ordinary Portland Cement (OPC) remains the key binder but to diminish its economical constraints, suitability of alternate binders has been examined worldwide. The present study aimed to investigate the effect of partial replacement of OPC with fly ash on the CPB’s strength and to determine the most optimal mix to achieve the required strength (1 MPa at 28 days of curing) at the most cost-effective way using the Bayesian network (BN). The CPB mixes were prepared at 72 wt.% solid concentration with mill tailings (87–91%), OPC (6–13%), and fly ash (0–4%), and instantly after mixing, fresh (slump, bleeding, density) CPB properties were measured. The strength was tested at 7, 14, 28, and 56 days of curing and initially analysed through traditional model. The traditional models follow the aleatory principle and are considered not appropriate for geotechnical engineering. Hence, the BN model was developed and tested. The reliability of two classifiers in learning model structure was compared which gives Naïve Bayes as the highest reliable tool. The CPB’s strength is most sensitive to the OPC content. The most consistent mix(s) is mill tailings: 87–88%, OPC: 9–11%, fly ash: 1–4%. Adding fly ash at 89–91wt% mill tailings possesses high failure probability of the CPB. The collinearity test indicates that the fines percentage and chemical composition of CPB’s ingredients are highly correlated with its slump, bleeding, and strength development.

Jul 14, 2022