Dewatering—the process of removing water from mineral slurries and concentrates—sits at the intersection of product quality, water stewardship, and operating cost in virtually every mineral processing operation. With few exceptions, the separation processes that recover valuable minerals rely on water as the carrier medium, and at the end of those circuits the concentrate must be separated from a pulp where water can constitute 60–80% of the total mass. Dewatering that concentrate to a shippable, storable, or directly saleable form is therefore non-optional, and doing it efficiently determines both the value realised from the product and the volume of water available for recycle. This guide provides a comprehensive treatment of all dewatering technologies deployed in modern mineral processing: thickeners of every type, vacuum and pressure filters, centrifuges, the flocculants that make sedimentation practical, water recycle circuit design, product moisture specifications, and the diagnostic framework for troubleshooting common dewatering problems.
Why Dewatering Matters
The strategic importance of dewatering extends well beyond the straightforward goal of drying a concentrate for shipment. Efficient dewatering is central to three interconnected plant objectives: water recovery and recycle, concentrate quality, and tailings management.
Water is increasingly recognised as a constrained resource in mining, particularly in arid regions where fresh-water allocation to industry is subject to regulatory caps and community scrutiny. A well-designed dewatering circuit recovers the large majority of process water for immediate reuse, dramatically reducing fresh-water demand. Modern thickeners and filter presses can collectively return upward of 90% of process water to the plant circuit, and paste thickeners designed specifically for high-density tailings can recover 70–80% of the water from tailings slurries before deposition (McLanahan, 2023).
Concentrate quality is directly tied to moisture content. Copper, lead, and zinc concentrates are now commonly ground to 80% passing 30 μm, and smelter penalty structures impose charges above 8–10% moisture by weight. At these fineness levels, achieving target moisture by vacuum filtration alone is difficult; pressure filtration has become the industry norm. Coal producers face similar constraints from rail-freight specifications and spontaneous combustion limits.
Tailings dewatering determines the footprint, stability, and closure risk of a tailings storage facility. Conventional slurry tailings with 30–40% solids require large impoundment volumes and complex water management. High-density, paste, and filtered dry-stack tailings represent progressive improvements in water recovery and geotechnical stability, all achievable through progressively more intensive dewatering upstream of the deposition point. The three dewatering stages—thickening, filtration, and thermal drying—form a cascading sequence in which each stage handles the output of the previous one, removing water in decreasing proportions at increasing unit cost (Wills and Finch, 2016).
Thickeners: Conventional, High-Rate, and Paste
Gravity sedimentation in a thickener is the most widely applied and most economical dewatering step in mineral processing. By allowing solids to settle under gravity and withdrawing the clarified overflow as process water for recycle, a thickener can remove up to 80% of the water from a dilute feed slurry, producing a thickened underflow of 55–65% solids by weight before any filtration is required.
Conventional Thickeners
The conventional thickener is a large, shallow, open cylindrical tank ranging from about 2 m to 200 m in diameter and 1–7 m in depth. Feed slurry enters through a central feedwell that is designed to dissipate momentum, de-aerate the slurry, dilute the feed if necessary to optimise flocculation, and distribute the settled material evenly. Clarified overflow discharges over a peripheral launder and is collected for recycle. Settled solids are swept to the central discharge cone by slowly rotating rake arms—the rake tip speed is typically about 8 m/min at the perimeter, corresponding to approximately 10 revolutions per hour for a 15 m diameter unit—and withdrawn as thickened underflow by positive displacement pumps. The low rotational speed means energy consumption is extremely low; a 60 m diameter thickener may require only a 10 kW drive motor (Wills and Finch, 2016).
Thickener tanks are manufactured from steel for diameters up to about 25 m and from concrete or a combination of steel and concrete for larger units. The rake mechanism is supported by a central superstructure (bridge or beam thickener) for units up to about 45 m diameter, or by a stationary center column for larger units. Traction thickeners, where a single long arm is driven by traction wheels running on a wall-top rail, are manufactured up to 200 m in diameter and are used in some of the world’s largest copper, iron-ore, and alumina operations.
High-Rate Thickeners
High-rate thickeners achieve a far greater clarifying capacity per unit of tank area than conventional thickeners by optimising flocculation in a specially designed feedwell. The feedwell of a high-rate unit induces controlled turbulence that promotes intimate contact between the flocculant and the solid particles, forming larger, faster-settling flocs. Because the area requirement of a thickener is governed by the need to keep the upward liquid velocity below the settling velocity of the slowest-settling particle, faster flocs allow a smaller tank to handle the same mass flow rate. In practice, high-rate thickeners can handle two to four times the throughput of a conventional thickener of equivalent diameter, which translates directly into capital cost savings on large projects.
High-Density and Paste Thickeners
High-density and paste thickeners extend the conventional thickener design with steeper cone angles and taller tank side walls. The additional bed height increases the compressive load on the sediment, squeezing out interstitial water and producing a denser underflow than is achievable in shallow conventional units. Paste thickeners, which are the most extreme form of this design, produce an underflow with a yield stress exceeding about 150 Pa—a material that is non-segregating and behaves rheologically as a visco-plastic rather than a Newtonian fluid. For a typical iron ore tailings, a conventional thickener might achieve 45–50% solids in the underflow; a paste thickener treating the same material can deliver 65–70% solids (McLanahan, 2023).
The non-Newtonian behaviour of paste underflows introduces a practical challenge: the torque required to rotate the rake arms in a dense, high-yield-stress bed is far greater than in conventional thickeners. Rake tip loads and torque ratings must be substantially higher, and an automatic lift mechanism is essential to raise the rakes when torque exceeds design limits. Phenomenon such as “rotating beds” or “islands” of solidified material that rotate with the rakes rather than being swept to the discharge must be monitored and managed through careful control of bed level and underflow pump rates (Wills and Finch, 2016).
Applications of paste thickeners include mine paste backfill preparation, where a high-solids, pumpable product is sent underground, and surface tailings disposal as paste or thickened discharge, both of which are discussed in detail in the tailings management guide in this series.
Thickener Sizing: The Coe-Clevenger and Talmage-Fitch Methods
Thickener area is determined by the need to limit upward liquid velocity below particle settling velocity at every dilution within the thickener. The two standard laboratory-scale sizing methods are both based on batch settling tests conducted in graduated cylinders:
The Coe and Clevenger method (1916) conducts multiple settling tests at a range of pulp dilutions spanning the feed to discharge composition range. For each dilution, the initial settling rate and the corresponding liquid-to-solids ratio are recorded. The required area is calculated at each dilution as a function of the mass flow rate of solids; the dilution giving the maximum area requirement is the critical dilution, and that maximum area governs the design with a safety factor of 1.2–1.5 applied.
The Talmage and Fitch method (1955) applies the Kynch sedimentation theory to a single settling curve, extracting the local settling rate at each point by constructing tangents to the curve. A simplified version of the method identifies the compression point on the settling curve and uses the tangent at that point to calculate the required area for the target underflow concentration. Both methods have documented limitations and must be used with empirical safety factors; pilot-scale testwork in a small replica thickener is the most reliable sizing basis for large installations (Wills and Finch, 2016).
Filtration: Vacuum and Pressure Filters
Filtration is the second dewatering stage in most mineral processing circuits, taking the thickener underflow from 55–65% solids to 80–90% solids (approximately 10–20% moisture by weight). The driving force for filtration is a pressure differential across the filter cake and medium: in vacuum filters that differential is approximately one atmosphere (about 0.1 MPa), while in pressure filters it can reach 6–8 bar or more. Higher pressure differential generally produces drier cake, which is why the industry trend is strongly toward pressure filtration for fine concentrates.
Pressure Filters
Pressure filters have become the dominant choice for copper, lead, zinc, and other fine concentrates requiring moisture below about 10% by weight. They come in two main configurations based on the direction of pressure actuation.
Horizontal plate (vertical frame) pressure filters suspend vertical polymer filter plates from a horizontal steel frame. Between adjacent plates hangs a filter cloth. The plates are hydraulically pressed together, slurry is pumped into each chamber, and filtration begins immediately as filtrate passes through the cloth. When the chamber fills with cake, a membrane squeeze is applied to compress the cake, followed by an air blow-through cycle at pressures up to 8 bar that achieves the final moisture reduction. The cake is discharged by gravity when the press opens. With cycle times of approximately 10 minutes and an upstream feed tank providing buffer storage, the operation appears continuous to the plant even though each individual press cycle is batch. Horizontal pressure filters can produce cake moistures of 8–10% on fine copper concentrates, eliminating the need for a thermal dryer in many operations (Wills and Finch, 2016).
Vertical pressure filters stack chambers vertically with a continuous filter cloth advancing through the stack. Operation is similar to horizontal units, but the continuous cloth advance allows more rapid cake discharge and is better suited to materials that are sticky or prone to cloth blinding.
For ultrafine materials finer than about 10 μm, where capillary pressures are very high and standard pressure filters cannot achieve adequate dewatering, the tube press is used. By enclosing the filtration in a tube and applying pressures up to 100 bar, the tube press can dewater kaolin, fine coal, and similar difficult materials to cake moistures that would be impossible with conventional equipment.
Vacuum Filters
Vacuum filters were the workhorse of mineral processing dewatering for most of the twentieth century and remain in widespread use for coarser, more free-draining materials and where pressure filtration economics are not justified.
The rotary drum filter is the most common vacuum filter type. A horizontal drum, partially submerged in a slurry trough, rotates slowly at 0.1–3 rpm. The drum surface is divided into compartments connected to a rotary vacuum valve; as each compartment rotates through the slurry zone, vacuum draws filtrate through the cloth and builds a cake on the drum surface. The cake is then dried as it rotates above the slurry, washed if required, and discharged by a knife, air blow, belt, or string mechanism. Drum filters can handle 50–300 kg of dry cake per square metre of filter area per hour, depending on material properties.
The ceramic disc filter represents a significant advance over conventional cloth disc filters. Microporous ceramic sectors exploit capillary action to draw liquid through very fine pores without requiring large vacuum pumps. Because the filter medium itself is hydrophilic and has fine pores, no air is drawn through when the filter is submerged in slurry, dramatically reducing vacuum pump energy consumption. Ceramic disc filters can approach the moisture performance of pressure filters on many mineral concentrates while consuming roughly half the energy of conventional vacuum disc filters.
The horizontal belt filter is used where water recovery rather than minimum cake moisture is the priority. Applications include tailings dewatering where space is limited and environmental restrictions prevent pond disposal, and recovery of valuable leach liquors in hydrometallurgical circuits. The endless perforated drainage deck supporting the filter cloth provides long contact time between slurry and vacuum, and the stage lengths (filtration, washing, drying) can be adjusted by varying the number of suction boxes.
Hyperbaric Filters
A hyperbaric filter places a conventional disc or drum vacuum filter inside a pressurised vessel, increasing the effective pressure drop to four bar or more. Cake moistures that would typically be 15% under vacuum can be reduced to approximately 8% under hyperbaric conditions, closing the gap with pressure filtration. Cake discharge from a pressurised vessel requires a rotary valve or lock-hopper mechanism, adding mechanical complexity.
Centrifuges in Mineral Processing
Centrifuges extend the principle of gravity sedimentation by replacing gravitational acceleration with the much larger centrifugal acceleration of a spinning bowl. The solid bowl (decanter) centrifuge is the most widely used type in the minerals industry, valued for its ability to handle feeds ranging from 0.5% to 70% solids at 0.25–100 t/h of dry solids, with particle sizes from 12 mm down to 2 μm.
In a decanter centrifuge, a horizontal cylindro-conical bowl rotates at 1,600–8,500 rpm; a screw conveyor inside the bowl rotates at a slightly different speed, continuously conveying settled solids up the conical beach and out through discharge ports at the narrow end. Liquid overflows ports at the wide end of the bowl. Moisture contents of 5–20% are achievable, depending on particle size distribution and the length of the conical beach. A long, shallow beach increases residence time in the draining zone and produces drier solids; a short, steep beach sacrifices moisture content for increased liquid clarity.
Centrifuges are more energy-intensive and mechanically complex than thickeners but considerably more compact. They are often the preferred solution for fine coal dewatering, where the need to avoid a thermal dryer on minus 0.5 mm material makes efficient mechanical dewatering critical. Flocculant use in centrifuges is limited because the high shear forces within the spinning bowl tend to break up flocs; the benefit of flocculation in improving settling is largely offset by the re-dispersion caused by the conveyor action.
Flocculants and Their Role
Flocculants are the key to making gravity sedimentation practical in mineral processing. Without them, fine particles of only a few micrometres diameter would settle so slowly in a thickener that commercially viable throughputs would be impossible.
Coagulation
Coagulation involves modification of the electrical double layer at the particle surface to reduce the zeta potential and allow van der Waals attractive forces to dominate. In practice this is achieved by adding multivalent inorganic cations (aluminium sulfate, ferric sulfate, ferric chloride, lime) or organic cationic polymers of low molecular weight. The concentration of multivalent cation required to reduce zeta potential to near zero is the critical coagulation concentration (CCC); trivalent ions such as Al3+ and Fe3+ are far more effective than Ca2+, which is in turn far more effective than Na+.
Flocculation
Flocculation uses high molecular weight polymers (typically over one million g/mol) to form molecular bridges between particles. The dominant commercial flocculants are polyacrylamide (PAM) copolymers, available in anionic, cationic, and non-ionic forms. Anionic polyacrylamides are most widely used in mineral processing because they tend to have the highest molecular weights and are least expensive. The bridging mechanism requires the polymer to be firmly bonded to one particle while presenting bonding sites to other particles; excess polymer can cause re-dispersion by saturating all available surface sites.
In practice, flocculants are prepared as dilute stock solutions (0.5–1%) and further diluted to a maximum of about 0.1% before addition. They must be added at carefully designed points in the feed stream with sufficient agitation to ensure good contact, but gentle enough conditions thereafter to avoid destroying the flocs. The age of the stock solution significantly affects performance: fresh solutions of high molecular weight PAM can lose effectiveness over time due to degradation of the polymer chains (Wills and Finch, 2016).
The relationship between flocculant optimised for thickening and flocculant optimised for filtration is not straightforward. Large, open flocs produced by high molecular weight PAM settle quickly and are ideal for thickening, but they trap water within their structure, which increases filter cake moisture. Lower molecular weight flocculants produce smaller, denser flocs that give better filtration performance. Where thickening feeds directly into filtration, the flocculant selection may require a compromise or a staged addition strategy.
Water Recovery and Recycle Circuits
Closing the plant water circuit—returning thickener overflow, filter press filtrate, and other process waters to the mill for reuse—is both an economic and environmental imperative. Fresh water costs money to purchase, treat, and pump; process water leaving the site requires treatment and carries regulatory obligations. A well-managed water balance can reduce fresh water consumption to a small percentage of total circuit water flow.
The primary source of recovered water in most plants is the thickener overflow. In a concentrate thickener, the overflow carries reagent residues from flotation that can interfere with performance if recirculated directly; the plant water balance and the tolerance of flotation for process water quality must be assessed carefully. In many operations, reagent residues in recycle water actually improve flotation performance once a steady state is established, but certain reagents, particularly those that froth strongly, can cause problems at elevated concentrations.
Tailings thickeners are the largest single source of recovered water in a concentrator. By thickening tailings to 55–65% solids before pumping to the tailings storage facility, the water returned to the mill in the thickener overflow can represent 60–70% of all process water entering the circuit. Paste thickeners and filter presses applied to tailings recover an even higher proportion, reducing the volume of water deposited with tailings and therefore the size of the reclaim pond required at the tailings facility.
Reclaim water from the tailings pond itself—recovered by floating pumps or fixed decant structures—is a further significant source. Managing the quality of this water is important: in sulfide-bearing tailings, contact with the stored solids enriches the water in heavy metals, sulfate, and acid; in cyanide operations, cyanide and its complexes can accumulate. Water quality analysis and appropriate treatment before reuse are essential for both plant performance and environmental compliance.
FLS (formerly FLSmidth) notes that modern thickening circuits integrated into closed-loop water management can significantly reduce freshwater intake while maintaining product quality and environmental standards (FLS, 2023).
Product Moisture Specifications
The moisture specification for a dewatered product depends on the downstream disposition of that product: shipment to a smelter, transport by rail or ship, storage in a stockpile, use as mine backfill, or disposal in a tailings facility.
For base-metal concentrates shipped to smelters, moisture is a paid-for impurity that increases freight costs and reduces the net effective grade of the shipment. Smelter contracts typically specify maximum moisture of 8–10% by weight for copper, lead, and zinc concentrates. Fine concentrates (80% < 30 μm) are more difficult to dewater and achieving these targets requires pressure filtration in most modern plants. Thermal drying to 3–5% moisture is used where pressure filtration alone cannot achieve the required specification, or where the concentrate must be stored in sealed containers to prevent dust generation.
Moisture affects the handling behaviour of stockpiled concentrates. Above the critical moisture content, fine concentrates become fluid under vibration or shock loading, a phenomenon called liquefaction, which has been responsible for fatal incidents aboard bulk carriers. Regulatory limits on the transportable moisture limit (TML) of mineral concentrates are specified by the International Maritime Organization (IMO) under the IMSBC Code, and producers must certify that shipped moisture is below the TML.
For mine backfill, the target is a pumpable paste with sufficient water to be transported by pipeline but enough solids content to develop adequate mechanical strength after curing. Cemented paste backfill is typically prepared at 70–80% solids by weight, which requires paste thickening of the tailings feed. The balance between water content and binder dosage is a critical optimisation variable because cement is the most expensive component of paste backfill.
Troubleshooting Common Dewatering Problems
Dewatering circuits are susceptible to a range of operational upsets, most of which can be diagnosed and corrected through systematic analysis of process data and physical sampling.
Thickener Overflow Turbidity
Cloudy or turbid thickener overflow is the most commonly encountered thickener problem. Causes include insufficient flocculant dosage, incorrect flocculant type, flocculant addition at a poorly designed point, changes in feed particle size or slurry chemistry, and bed level exceeding the thickener capacity. The first diagnostic step is a flocculant dosage trial: if turbidity clears at higher dosage, the solution is straightforward. If not, flocculant type, stock solution age, and addition point should be systematically investigated. A bed level that is too high can cause fine particles to short-circuit to the overflow; reducing underflow pump rate is not always the solution, as it may worsen the problem by allowing the bed to grow further.
Thickener Underflow Too Dilute
If the underflow is thinner than target, the bed level may be too low (insufficient residence time for compaction), the underflow pump rate may be too high, or the rake torque may be insufficient to transport solids to the discharge cone. Bed level monitoring using ultrasonic, conductivity, or pressure-based sensors is essential for control; operators without reliable bed level indication are flying blind. High rake torque before the bed level reaches design limits suggests poor bed structure, possibly caused by over-flocculanting or by physical disturbance of the bed by excessive turbulence in the feedwell.
Filter Cake Moisture Too High
Excessive filter cake moisture can result from: too-fine particle size distribution in the feed (capillary pressures exceed applied vacuum or pressure); insufficient filtration pressure (check system pressure, membrane condition, and cloth integrity); cloth blinding by fine particles or precipitates (regular cloth washing and periodic replacement required); and inadequate drying time (adjust cycle duration or reduce belt speed). Adding a lower molecular weight flocculant to the thickener underflow before filtration can improve cake permeability by producing a more open, uniform cake structure.
Pump Blockages in Underflow Lines
High-density underflow, particularly from paste thickeners, is prone to blockages in the underflow piping, especially at bends, valves, and where piping slopes upward. The underflow lines must be as short and straight as possible. Underflow pumps should never be stopped without first diverting to recirculation back into the feedwell; a static column of high-yield-stress paste will solidify in the pipe within minutes if flow is interrupted. Variable-speed underflow pumps controlled by a nuclear density gauge on the underflow line maintain a consistent underflow density and minimise blockage risk (Wills and Finch, 2016).
Flocculant Ineffectiveness
Flocculant performance can deteriorate without an obvious cause. Common reasons include: stock solution degraded by excessive age or storage temperature; changes in process water chemistry (pH, ionic strength, dissolved organics from flotation reagents); and changes in ore mineralogy affecting the surface chemistry of the particles. Bench-scale jar tests using current process water and current ore samples should be repeated whenever thickener performance changes unexpectedly; the optimal flocculant type and dosage can shift significantly between ore domains.
References and Further Reading
- Wills, B.A. & Finch, J.A. (2016). Wills’ Mineral Processing Technology, 8th Edition. Butterworth-Heinemann/Elsevier.
- Engineering & Mining Journal. (2015). Dewatering: An Increasingly Important Mineral Process. E&MJ.
- Encyclopaedia Britannica. Mineral Processing—Dewatering, Filtration, Separation.
- FLS (2023). Thickening and Water Stewardship in Mining. FLSmidth.
- McLanahan. (2023). Paste Thickeners. McLanahan Corporation.
- McLanahan. (2023). The Role of Thickeners and Filter Presses in Tailings Management.
- Diemme Filtration. (2023). High-Rate Thickeners in Mineral Processing. Diemme Filtration.