Search Documents

By A Cheatle

GoldcorpÆs Musselwhite mine in north-western Ontario, Canada has in recent years become a world-class deposit. Initially discovered in the early 1960s by Alan and Harold Musselwhite, the then marginal mine eventually started production in 1997 and has to date produced 2.6 Moz of gold. Two million ounces of gold are currently classified as Mineral Reserve and a further 1.2 Moz of gold are classified in the Mineral Resource category for a total gold inventory approaching 6 Moz. Growth of the mineral inventory has been achieved through the successful implementation of strategic planning that sought to extend the mine life through discovery of additional economic zones and increased mined tonnages. Accelerated exploration activity commenced in 2001, at a time when gold price was at its lowest in decades. Focus was placed on building, developing and retaining a highly skilled geological team, maintaining purpose, conducting external reviews and developing a new geological model through research partnerships. Ongoing collaborative research between Lakehead University and Musselwhite mine has seen the development of new structural and geological models through post-graduate and academic studies. Research has included lithogeochemical, stable isotope and structural and metamorphic studies and has led to the evolution of a new mineralisation model (for the mine) and the recognition of brittle and ductile deformation regimes within the mineralised horizons. Exploration drilling continues to provide support for the new geologic model. Gold discoveries have extended known mineralisation 2.3 km along strike from mineral reserves and laterally into parallel zones. The exploration potential to further increase production rates, Mineral Reserves and Resources remains very positive and increases in overall gold inventory are likely. Musselwhite mine has significantly exceeded initial projections for a ten year life-of-mine and a marginal existence.

Jan 1, 2009

By Morris Steinberg, John Wulff, Robert Kamm

IN powder metallurgy it has often been observed that shrinkage may occur in one direction and growth in another during sintering. Even in long-time sintering experiments the rate of shrinkage may be different in different parts of the compact. In the fabrication of some parts requiring a constant permeability it is even questionable in some cases whether pressing from one direction should be attempted. These observations require for their interpretation a more detailed study of green or pressed density than has heretofore been available. Usually average density measurements of the whole compact are made and in some cases carefully cut sections are measured. A different approach to the problem was first mentioned by Rakowski1 and also used by Balshin.2 Balshin's diagram is reproduced in Fig r and schematically shows the result of pressing three layers of metal powder with fine intervening layers of graphite. The retarding effect of the die walls and a qualitative example of the density distribution is evident. More quantitative information about the distribution of density within a cold-pressed powder compact would be useful in showing the variation in density in compacts as influenced by the type of powder, shape of die, lubrication during the process, and other variables. Furthermore, information about the cracking of cold-pressed compacts may be obtained from an analysis of the stress distribution within a powder compact, and some information about the stress distribution can be obtained in turn from the density and strain distributions. A more detailed report on some of the present experiments appears in reference 3. PRESENT INVESTIGATION In order to ascertain the density gradients in cold-pressed metal powders, a deformable lead grid within the powder was employed. After pressing in cylindrical dies the ejected compact was radiographed, giving a pattern of the deformed grid. From the radiograph density and deformation from point to point is readily measured. From the radiograph strain distribution and stress trajectories were also determined. The latter also permits the determination of the coefficient of friction at the die wall. This technique indicates that the densest part of a compact pressed from one side is at the top outer circumference and the least dense region at the bottom circumference near the stationary plunger. For some heights of compacts the density at the axis is such that the lower density is at the top rather than at the bottom. So irregular a density distribution is due primarily to die-wall friction. As the ratio of diameter to height of compact is increased the die walls play a correspond¬ingly smaller role. More uniform density

Jan 1, 1947

By Margaret R. Lickes

Mercury was produced originally at the Idaho Almaden mine from 1939 to 1942. When the ;.nine was shut down, more than 100,000 tons of ore, containing from 2.5 to 3 pounds of mercury per ton, was known to be adjacent to old workings but had been considered too low grade for commercial operation. Later, detailed, systematic exploration by Rare Metals Corporation of America increased and delimited enough minable reserves for further operation to be foreseen for several years. The average grade of the ore is estimated at slightly more than 3.5 pounds of mercury per ton. Open-pit mining and installation of the largest rotary kiln in the United States furnacing mercury ore have made possible profitable exploitation of these reserves (fig. 1). The bulk of the exploration was by percussion drilling, although some diamond drilling was done to check the results. When enough reserves were indicated to assure a successful operation, stripping and plant construction were started simultaneously. Figure 2 is a general view of the plant. The ore bodies, although irregular in size, shape, and grade and covered by essentially barren cap rock, are amenable to low-cost, trackless, open-pit mining. Development consists of stripping the overburden and removing, barren cap rock. Ore is being mined at approximately [5],000 tons per month and treated in a 175-ton-per-day furnacing plant.

Jan 1, 1957

By E. J. Garrison

A.C. Noble presents a clear and concise summary of the factors pertinent to a competent reserve evaluation of a mineral occurrence. However, I believe that he confuses the difference between a resource and a reserve. By definition, a reserve estimate includes all known economic, legal, mining, metallurgical and environmental factors impacting the recovery of the product(s) found in the deposit. A resource estimate, however, is an estimate of what is there - what is available for exploitation and should not include economic and other considerations. Only after defining the resource, its extent, location, physical and chemical properties, can an intelligent selection process begin of the extraction and processing technologies most appropriate to the circumstances encountered. The early and almost certainly inappropriate application of these often cookbook choices has probably lead many companies to walk away from good properties. (How many mines have been found in others rejects?) An example of the process is the early selection of metallurgical samples, after encouraging values are found in a few holes or surface samples, in an attempt to determine the economics of the prospect. Unfortunately, the sampler does not know how representative of the occurrence the sample is because he does not know what the deposit looks like. As a consequence, the property becomes saddled with a bad metallurgical sample that is often inappropriately used in evaluating the deposit which it does not represent.

Jan 1, 1994

By Abdelaziz Baçaoui, Imane Aarab, Abdelrani Yaacoubi, Khalid El Amari, Abderahman Etahiri

Phosphate ores upgraded by flotation are widely used as vital raw materials for the sustainable production of phosphate fertilizers to maintain the food supply. High-grade phosphate reserves become more and more depleted. Low-grade phosphate mining and processing are in the spotlight of the fertilizer industries to cope with the exponential growth in the world’s population and meet the increased demand for fertilizers. As part of a series of research works conducted within a major project aiming at upgrading a low-grade phosphate ore never-exploited before, the present work endeavors to optimize the flotation of this ore using the response surface method (RSM) as a follow-up of a screening study. Unveiling new insights on such a strategy paved the way for improving flotation performances. This work demonstrated that an optimum 30.18% P2O5 with 88.07% recovery was achieved using 63.10 g/t of sodium alginate (NaAL) and 2484.72 g/t of sodium oleate (NaOL) at pH 11.40. This maximum was contrasted with a new formulation of fatty acids (FAM). The 30.76% P2O5 grade with a recovery of 89.72% was attained using 50.94 g/t of NaAL and 1455.40 g/t of FAM at pH 11.68. FAM collector markedly outperformed NaOL at low concentrations proving the efficiency of synergists.

Dec 7, 2022

By D. E. Wolfson, D. A. Reynolds, F. W. Smith

IN 1948 the U. S. Bureau of Mines began a three-phase program to evaluate the extent and quality of U. S. coking coal: 1) a factual appraisal of known recoverable reserves in beds of mineable thickness, 2) a study of preparation characteristics, and 3) a study of carbonization properties. A previous paper presented the methods used in assessing reserves.' Although this investigation is only partly completed, data on coking properties of the Pittsburgh district coals are now available. The results are presented in this article. When coals are selected for manufacture of metallurgical coke more factors must be considered than for any other major use, as no basic industry requires coal of such high quality and uniformity. This is particularly true in the Pittsburgh district, where blast furnace practice is based on use of cokes from high quality coals. Although it is true that inferior coals are carbonized in other industrial areas, both here and abroad, the Pittsburgh coals are the best economically obtainable in the required tonnages, even though they are not the most desirable for metallurgical use. In the general evaluation of suitability of coals for coke making, particular attention is given to: 1) reserve, 2) ash and sulfur content, 3) inherent properties of forming strong, well fused coke, 4) expanding characteristics, and 5) blending properties. Items 1 and 2 are covered in detail in the first two phases of the survey, and coking behavior is considered in items 3 through 5. Problems is Coal Evaluation: Although not all coal beds of immediate or potential use to the coking industry in the Pittsburgh area have been tested for their coking properties, enough tests have been made to permit certain generalized statements. Virtually all the coals that will be used in this district are from the Appalachian region, and except for high oxygen coals from outcrops or stripping operations, they are coking coals. Hence the evaluation problem is not so much that of distinguishing between coking and noncoking coals as attempting to measure quantitatively differences in the coking and expanding properties of the better coking coals. Quantitative differentiation between good coking coals, is complex, since differences in their coking abilities are frequently of about the same magnitude as the precision of the test. Small-scale tests are generally unsuitable for measuring these slight differences, and there has been gradual acceptance of pilot-scale tests. Test Methods Employed: All the data reported here were obtained in two units known as the Bureau of Mines-American Gas Association (BM-AGA) carbonization apparatus and the sole-heated expansion oven. The BM-AGA unit, which consists essentially of a cylindrical steel retort about 18 in. diam and 26 in. high, carbonizes enough coal (180 lb) to allow evaluation of the coke by standard methods of the American Society for Testing Materials. The BM-AGA carbonization procedure and methods of testing the gas, liquid, and solid products have been described.' Expanding characteristics of the coals were determined in the sole-heated expansion oven," wherein the coal charge, heated from the bottom only, is carbonized under a constant applied force and the linear expansion or contraction is measured. Although results in this oven have never been quantitatively correlated with the expansion behavior of coals in conventional slot ovens, they are useful in assessing expansion behavior. In this report discussion will be confined to results obtained from these pilot-scale units, although the following data are available for most coals: proximate and ultimate analyses, free swelling index, agglutinating value, and plasticity. Sources of Coals Tested: In selection of coals for coke manufacture, economic considerations are as important as quality, and each coke plant balances these two factors to meet requirements of the most satisfactory coke at lowest price. Although these economic factors may in some instances outbalance small differences in coke quality, they are usually specific to each coke plant and cannot be given in any general statement. Recognizing that coal and plant operating costs are beyond the scope of the survey, the USBM evaluates the coking quality of various coals without regard to their competitive economic status. For the purpose of this article, coals of interest to the Pittsburgh district are arbitrarily defined as

Jan 1, 1958

By S. F. Mauser

This report describes the study of factors which affect medium voltage (661 - 1000 volts) trailing cable splices used in underground coal mines. Improvements in quality, dependability, and design of splice kits and shop splices are presented in the study results and recommended guidelines and regulations. Recommended research programs are included which describe approaches for development of the recommended guidelines and regulations for medium voltage trailing cable splices. The basis for the conclusions and recommendations presented in this report is included in three appendices which describe a survey, splice examination, and general requirements development prepared during the progress of this study.

Jan 1, 1977

By John Dasher

The analyses reported on in this paper indicate that there are hydrometallurgical routes from sulphide concentrates to copper that are potentially competitive with smelters which are not permitted to emit sulphur oxides. The various hydrometallurgical processes, and their economics, are discussed.

Jan 1, 1973

By N. E. Maccallum

(Cleveland Meeting, October, 1912.) ALMOST half a century has passed since the Siemens brothers, after tedious and costly experiments, finally began the manufacture of open-hearth steel. The furnace of that time was very small, having a hearth-capacity of from 3 to 4 tons, and we can appreciate the wonderful development that has taken place in those 50 years by the knowledge that to-day furnaces of 50 times the original capacity are in operation. As the new steel process assumed importance, and furnaces began to grow in number and in size, a question arose, to which no satisfactory answer has ever been given, and that was: ?What is the safe and economical limit of the open-hearth furnace?" Various stopping-places have been suggested, and 20, 30, 40, and 50 tons have each been advocated, but all such arbitrary restrictions have been passed, and 60 tons would probably represent the average furnace, although in a few instances from 70 to 80 tons are regularly tapped into one ladle. The furnaces which are the subject of the present paper have gone far beyond even these limits, and since, in my opinion, they mark another step in the progress of open-hearth steel, they will be considered fairly and frankly, in order that those to whom the matter may be of interest may be able to judge and form their own Conclusions. Short accounts of the work of these furnaces have appeared from time to time in various technical journals, of which the following is a brief résumé. These furnaces are the property of the Phoenix Iron Co., of Phoenixville, Pa., and their distinctive and important feature is, that they charge to twice the capacity of one ladle, and tap simultaneously into two ladles, with a device for controlling the flow into either. In this way the capacity of the furnace is not limited, as formerly, but heats of any size within the capacity of two ladles may be charged.

Oct 1, 1912

By R. O. Dannenberg

The Bureau of Mines conducted laboratory investigations with the objective of developing an alternative process for recovering cobalt and copper from complex sulfide concentrates. Processes currently used depend largely upon precipitation techniques to remove impurities and recover cobalt from sulfate liquors prepared by oxidative pressure leaching of the sulfide concentrates. The Bureau conducted bench-scale research on a process for treating cobaltite concentrates, comprising (1) oxidative pressure leaching, (2) jarosite precipitation followed by H202 oxidation and pH control to remove iron and arsenic, (3) copper solvent extraction with a mixed hydroxyoxime-amine extractant, (4) copper electrowinning from recirculating acidic strip liquor, (5) selective cobalt extraction from copper solvent extraction raffinate with a phosphinic acid extractant, and (6) electrowinning of cobalt from a recirculating weak acid strip liquor. Overall cobalt and copper recoveries were 91.7 and 84.1 pct, respectively. Electrowon products assayed 99.8 pct Co and 99.89 pct Cu.

Jan 1, 1987

By A. J. Basu

Geotechnical design plays a key role in mine design. I A proper geotechnical design provides a safe working environment and facilitates efficient mining operation. There are mainly three design approaches I) judgmental, based on observational and experiential facts, 2) empirical, based on empirical design formulae, and 3) analytical, incorporating numerical analysis. Most design applications are based on the design methods one and two as mentioned earlier. However, complexities arise in deep underground mining operations and deep open pit mines. These complexities and difficult geological environment introduce more uncertainties and simple empirical formulae would not be adequate as a design tool. These situations require numerical modelling and analysis of the design problems including interpretation of the results obtained from the analysis. There are several tools available for performing the numerical modelling and analysis. A brief review of the available tools was addressed earlier (Basu and Kuruppu, 1995). The authors selected ANSYS (a general purpose finite element analysis tool) for use in mining geomechanics and geotechnical design. This paper includes several examples demonstrating the applicability of this software in designing structures applicable to mining geomechnics. The authors hope that the design examples will demonstrate the use of the numerical modelling and analysis application in excavation design.

Jan 1, 1996

By B. W. Gandrud

In 1927 Coghill showed that there is considerable liberation of the iron oxide mineral from gangue when the Alabama red iron ores are ground to a size suitable for classification and tabling but that there is not sufficient to make possible a high extraction and a high grade of concentrate. Coghill states: High extraction and 'high grade of concentrate cannot be obtained without grinding to a size comparable with flotation or cyanide practice. If grade of concentrate be sacrificed the grinding need not be so severe. These ores slime readily during grinding, and in most instances the slimes are only slightly higher grade than the original ore. Not only do they carry a considerable percentage of the total iron in the ore, so that to reject them as tailings would result in prohibitively low extraction, but they are too fine to be amenable to ordinary methods of gravity concentration. In 1929 the Bureau of Mines published results showing that certain Alabama oolitic iron ores could be concentrated by classification and tabling after grinding, with rubber-covered rods. This selective grinding produced slimes of much higher grade than the original ore, and they were considered part of the final concentrate.

Jan 1, 1934

By F M. Senda

This paper highlights strategies used at the Barrick operated Kibali Gold Mine (KGM) to improve the drilling accuracy and blasting effectiveness for longhole open stoping. The mine currently uses Solo DL421 for the required stope drilling and has experienced an average deviation of 10 per cent. Strategies are raised to eliminate or at least mitigate the adverse effects of blasthole deviation and the impact of blasting induced effect at the stope faces including the stope crown. The integration of deviation trends in the design would allow the actual blastholes to be established within their required positions to then facilitate a proper energy distribution with the eventual production blasting. The implementation of stope undercutting reduces the hole length from 35 m to 30 m which in return decreases blasthole deviation by 3 per cent. Another strategy adopted to improve the drilling accuracy is to slow down the penetration rate with the use of several operating approaches, including 135 MPa percussion pressure, 65 rev/sec rotation speed and 17 MPa feed pressure. To this penetration rate, a constant k factor is applied to increase the rate when the bedding inclination is less than 45° from the horizontal or when the hole length is under 10 m. This reduces deviation by a further 4 per cent. Where the rock mass is highly jointed it is advised to use the average penetration rate over the entire length of the hole. Furthermore, the paper shows the effect of deviation on the blasting results through a prediction of fragmentation blasting model optimised with reduced deviation and highlights the effect of presplit mining and timing delay on the blasting effectiveness. Although the size pattern increases with deviation, a strategy pertaining to timing delay with blasting would be to align to the work of Grant (1990) where the optimum hole to hole delay is found to be between 7–10 ms/m of spacing, while the optimum delay between adjacent ring varied between 110–160 ms. Stabilisation of crowns and wings will perform with the implementation of presplit blasting.

Mar 29, 2023

By W. H. McAllister

The paper considers the geology and the coalfields that comprise the Scottish south area. It traces the history of known sources of spontaneous combustion and indicates the collieries concerned, with the position of the seams in the geological series where heatings have occurred. The methods of work over the years are described. It relates in detail the circumstances surrounding a fire at Bilston Glen Colliery in the Lothian Coalfield, and the heating at Killoch Colliery in the Ayrshire Coalfield, and the steps taken to deal with them. It also describes some of the difficulties in steep seam working. Methods adopted to prevent spontaneous combustion throughout the area are discussed, and some detail of the UNOR gas analyser is given. Finally, it comments on the present techniques in use in the area, particularly at Bilston Glen Colliery, to prevent spontaneous combustion.

May 23, 1905

By E. C. Lane

[The motor gasoline that s being marketed in the United States this summer slightly more volatile then that sold a year ago. This increase in volatility is shown by a general lowering ref the average distillation rang, of about 2°F. During the post 10 years the Bureau of Mines has made semiannual surveys of the gasoline marketed in the United States, the survey just completed being the twentieth in the series. The 12 cities in which samples were collected have beer chosen as representative of the more important marketing territories, as it is obviously impracticable to obtain a sample of every gasoline sold in the entire country.]

Jan 1, 1929

By L. Alan Weakly

The U.S. Department of Energy (DOE) has entered into a $l.5 million Cooperative Agreement with Alpine Equipment Corporation/Astro International Corporation ("Alpine") of State College, Pennsylvania, for the evaluation of water-jet-assisted (WJA) mechanical cutting of oil shale under actual mining conditions. A commercial-sized, heavy WJA mechanical miner is being assembled based upon the successful "factory testing" of a smaller prototype miner utilizing standard tungsten-carbide and Polycrystalline Diamond Composite (POC) bits in large oil shale boulders. Actual field testing of the machine will be conducted in an oil shale mine located in the Piceance Basin of Western Colorado. In-mine logistical support is being provided by several oil shale companies through the Colorado Mining Association (CMA) and the U.S. Bureau of Mines (USBM).

Jan 1, 1990

A list of permissible mine equipment approved by the Bureau of Mines before January 1, 1937, was published in Bureau of Mines Information Circular 6942.3/ Information Circular 7007,4/ published as a supplement thereto, listed equipment approved. Burin', the calendar year 1937. The present list includes most of the material liven in these two publications, as well as approvals that have been issued up to January l, 1940. Some equipment formerly listed has been omitted because it has become obsolete or is no longer actively marketed. The list now includes electrically-operated air compressors, coal drills, loading machines and conveyors, mining machine, mine ones, rock-dusting machines, room hoists, storage-battery gathering and haulage locomotives and storage-battery power trucks, mine telephones, junction boxes, electric cap lamps, flame safety lamps, electric hand and trip lamps, electric flood lens, flashlights, methane detectors, and single and multiple-shot blasting units. The appendix at the end of this circular lists trailing, cables that have passed tests prescribed by Schedule 2D. When equipment passes the inspections and tests prescribed by the United Status Department of the interior, Bureau of Mines, in published "schedules" or standards, it is formally approved as "permissible" for use in gassy mines, and the manufacturer is then permitted to mark it with an approval plate that identifies it as having been approved by the Bureau. The inspections and tests are designed to insure that each equipment satisfies the minimum requirements for safety in use. The Bureau's purpose in investigating equipment of this type and listing that

Jan 1, 1940

By J. Spotts McDowell

Preparation and Use.-Magnesite is an important refractory in open-hearth, heating, and electric furnaces for steel-making and in many of those employed in the metallurgy of copper and lead. It is sold in the form of brick, finely ground "furnace magnesite" for brick-laying, and dead-burned grains for making and repairing furnace bottoms. The latter are a mixture of granules varying in size from pieces of about 5/8 in. (1.6 cm.) diameter to very fine but sandy particles. Dead-burned magnesite results from calcining 'the crude or lightly burned mineral at a temperature that will not merely drive off practically all the CO2, but will also cause sintering of the particles. During this process the pieces shrink considerably and become hard, dense, and inert to atmospheric moisture and C02; under-burned material, on the other hand, will hydrate on exposure to the air. A small percentage of ferric oxide seems to be necessary for the production of a satisfactory sinter; from 4.5 to 8 per cent. in the dead-burned grains is considered the most desirable amount. Dolomite has been little used for brick-making in the United States, but it is prepared for use in the granular condition calcined or "double-burned" and is the principal ingredient of several materials offered for sale under various trade names for refractory purposes. Dolomitic refractories are almost wholly confined to the open-hearth and electric furnaces, where they are used for fettling and as substitutes for magnesite. Much more magnesite and dolomite are used for basic open-hearth steel-making than for all other refractory purposes. The hearth of the furnace is usually built up of magnesite brick and dead-burned grain magnesite so laid that the brick base is protected by a working bottom of the granular material. The latter is sintered into place in layers 1 to 2 in. (2.5 to 5 cm.) thick to a total depth of 12 to 18 in. (30 to 45 cm.) at the center of the furnace. After each heat, burned dolomite is thrown against the banks as high as it will stick, and all holes in the bottom are filled. At most plants such holes, at the end of each week, are also

Jan 2, 1919

By B L. Sainsbury, M F. Lee, A Penney

For most practical rock engineering purposes, stress versus strength conditions around underground openings for which rock masses fail and their various modes of failure, are reasonably well-understood and predictable. Some rock masses and particular rock types fail slowly and relatively quietly. However, others can fail rapidly, violently and maybe with the ejection of significant broken rock. In some countries and in some mines, the latter style of violent fracturing and failure of the rock mass is referred to as ‘strain bursting’. It can be alarming and extremely hazardous to nearby miners and machines. For such situations, the design and installation of appropriate ground support, to provide a safe work place, can be challenging. At any particular mine site, it’s often generally understood which local rock masses (rock types) are prone to strain bursting. Competent and strong siliceous rocks and some massive sulphides are often the main offenders. While some neighbouring, equally-competent and strong rocks might also fail, they do so more quietly and without ejection. To understand this contrast, a better understanding of the fracturing process is required other than simple and standard strength versus stress considerations. Suites of standard intact rock properties from numerous Australian mines have been used to define and help understand different high-stress failure styles for component intact rock; especially those prone to violent fracturing (strain bursting). Simple considerations of micro-fracturing (defects in the intact material), plus the available energy at failure versus the energy consumed during fracturing, distinguishes between over-stressed and competent rock types that fracture violently and those that don’t. CITATION:Lee, M F, Penney, A and Sainsbury, B L, 2018. Proneness of competent over-stressed intact rock to violent fracturing, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 170–184 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

By C. Kathryn Sole

The close proximity of cobalt and nickel on the Periodic Table and their frequent occurrence together in nature means that their separation in primary hydrometallurgical flowsheets is not only often necessary, but requires the use of innovative chemistry to achieve high selectivity. Today, increasingly stringent demands for the high-purity cobalt and nickel salts required for battery applications in consumer electronics, electric vehicles, and solar applications are again challenging the state of the art. The evolution of cobalt–nickel separations is traced from the basic precipitation technologies employed in 1960s, through commissioning of the first cobalt solvent-extraction plant in South Africa in 1974, to the commercialisation of organophosphinic-acid extractants (which provided a step-change improvement in selectivity) and their eventual widespread use in the 1980s and ’90s. This was followed later by their sulphur analogue, which enabled different possibilities for laterite processing in the new millennium. Separations in chloride media, which take advantage of the different speciation chemistries, and nickel extraction by oximes in ammoniacal systems are also discussed. More recently, chelating and solvent-impregnated resins have added further dimensions to the possibilities for flowsheet design and enabled higher purity products to be attained from both primary and secondary sources. Selected case studies and flowsheets are described.